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6-K

New Pacific Metals Corp (NEWP)

6-K 2026-08-14 For: 2026-08-13
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Added on August 14, 2026

UNITED STATESSECURITIES AND EXCHANGE COMMISSION****Washington, D.C. 20549

FORM 6-K

REPORT OF FOREIGN PRIVATE ISSUER

PURSUANT TO RULE 13a-16 OR 15d-16 OFTHE SECURITIES EXCHANGE ACT OF 1934

For the month of: August, 2026

Commission File No. 001-40381

NEW PACIFICMETALS CORP.

(Translation of registrant's name into English)

Suite 1750 - 1066 W. Hastings Street

Vancouver BC, Canada V6E 3X1

(Address of principal executive office)

[Indicate by check mark whether the registrant files or will file annual reports under cover of Form 20-F or Form 40-F]

Form 20-F ¨ Form 40-F x

EXHIBIT INDEX

EXHIBIT DESCRIPTION OF EXHIBIT
99.1 Carangas Project: NI 43-101 Technical Report and Preliminary Economic Assessment, Oruro, Bolivia

SIGNATURE

Pursuant to the requirements of the Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned, thereunto duly authorized.

Dated: August 14, 2026 NEW PACIFIC METALS CORP.
/s/ Jonathan Hoyles
Jonathan Hoyles
General Counsel & Corporate Secretary

Exhibit 99.1

CERTIFICATE OF QUALIFIEDPERSON

Kevin Murray , P.Eng.

I, Kevin Murray, P.Eng., certify that:

1. I am employed as a Principal Process Engineer with Ausenco Engineering Canada ULC, (Ausenco), with an<br>office address of 1050 West Pender, Suite 1200, Vancouver, BC, V6E 3S7.
2. This certificate applies to the technical report titled “Carangas Project: NI 43-101 Technical Report<br>and Preliminary Economic Assessment, Oruro, Bolivia” that has an effective date July 16, 2026, and report date of August 13,<br>2026 (the “Technical Report”).
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3. I graduated from University of New Brunswick with a Bachelor of Science in Chemical Engineering in 1995.
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4. I am a member in good standing of Engineers and Geoscientists British Columbia (No. 32350), Northwest<br>Territories Association of Professional Engineers and Geoscientists (No. L4940) and Association of Professional Engineers and Geoscientists<br>of Saskatchewan (No. 82404).
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5. I have practiced my profession continuously for<br>25 years. I have been directly involved in all levels of engineering studies from preliminary economic assessments (PEAs) to feasibility<br>studies. I have led preliminary test work design, test work analysis and flowsheet development as well involvement in detailed design<br>and commissioning. I have also developed operating cost estimates and contributed to and reviewed capital cost estimates. I have been<br>involved with gold leach studies including Argonaut’s Magino and Seabridge Gold’s Courageous Lake projects as well as lead-zinc<br>flotation studies including Prairie Creek.
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6. I have read the definition of “Qualified Person” set out in the National Instrument 43-101<br>Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional<br>association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections<br>of the Technical Report that I am responsible for preparing.
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7. I have not visited the project site.
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8. I am responsible for Sections 1.1–1.2, 1.15, 1.16.1, 1.17, 1.19, 1.20, 1.21, 2.1–2.3, 2.4.5,<br>2.5-2.7, 3.1, 17, 18.1–18.3, 18.7–18.8, 18.10, 19, 21.1, 21.2.1–21.2.2, 21.2.4–21.2.10, 21.3.1–21.3.2, 21.3.4–21.3.7,<br>22, 23, 24, 25.1, 25.10–25.12, 25.14–25.16, 25.17.1, 25.17.2.7–25.17.2.8.1, 25.17.3.6, 26.1, 26.6, 26.7.5, and 27 of<br>this Technical Report.
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9. I am independent of New Pacific Metals Corp. as independence is defined in Section 1.5 of NI 43-101.
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10. I have not been previously involved with the Carangas Project.
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11. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared<br>in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, information and belief,<br>the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be<br>disclosed to make those sections of the Technical Report not misleading.
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Dated: August 13, 2026

“Signed and sealed”

Kevin Murray, P.Eng.

Page 1 of 1

CERTIFICATE OF QUALIFIED PERSONScott C. Elfen, PE, P.Eng.

I, Scott C. Elfen, PE, P.Eng., certify that:

1. I am m employed as the Global Lead Geotechnical and Civil Services of Ausenco Engineering Canada ULC (Ausenco),<br>with an office address of 1050 West Pender Street, Suite 1200, Vancouver, BC, V6E 3S7, Canada.
2. This certificate applies to the technical report titled “Carangas Project: NI 43-101 Technical Report<br>and Preliminary Economic Assessment, Oruro, Bolivia” that has an effective date July 16, 2026, and report date of August 13,<br>2026 (the “Technical Report”).
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3. I graduated from the University of California, Davis, CA, in 1991 with Bachelor of Science degree in Civil<br>Engineering (Geotechnical).
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4. I am a professional engineer (civil) registered with the State of California in the United States of America<br>(No. C56527) by exam since 1996. I am a professional engineer registered within the province of British Columbia, Canada (No. 64064)<br>by Exam since 2025.
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5. I have practiced my profession continuously for 32 years with experience in the development, design, construction<br>and operations of mine waste storage facilities, such as waste rock storage facilities and tailings storage facilities ranging from slurry<br>to dry stack facilities, focusing on precious and base metals, both domestic and international. I have developed geotechnical design parameters<br>for pit slope design, plant foundation design, and other supporting infrastructure. Examples of projects I have worked on include Skeena’s<br>Eskay Creek Project PEA, PFS, and FS, O3 Mining’s Marban Project PEA and PFS, and First Mining Gold’s Springpole PEA and PFS.
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6. I have read the definition of “Qualified Person” set out in the National Instrument 43-101<br>Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional<br>association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections<br>of the Technical Report that I am responsible for preparing.
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7. I visited the project site on March 16 and March 17, 2026.
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8. I am responsible for Sections 1.16.2, 2.4.2, 12.3, 18.4, 18.5, 18.6, 18.9, 25.17.2.8.2, 26.5.2, 26.7.1-26.7.2,<br>26.7.4 and 27 of this Technical Report.
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9. I am independent of New Pacific Metals Corp. as independence is defined in Section 1.5 of NI 43-101.
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10. I have not been previously involved with the Carangas Project.
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11. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared<br>in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, information and belief,<br>the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be<br>disclosed to make those sections of the Technical Report not misleading.
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Dated: August 13, 2026

“Signed and sealed”

Scott C. Elfen, P.Eng.

CERTIFICATE OF QUALIFIEDPERSON

James Millard, P.Geo.

I, James Millard, P.Geo., certify that:

1. I am employed as a Director, Strategic Projects with Ausenco Sustainability Inc (Ausenco), with an office<br>address of 18-4515 Central Blvd, Burnaby BC V5H 0C6, Canada.
2. This certificate applies to the technical report titled “Carangas Project: NI 43-101 Technical Report<br>and Preliminary Economic Assessment, Oruro, Bolivia” that has an effective date July 16, 2026, and report date of August 13,<br>2026 (the “Technical Report”).
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3. I graduated from Brock University in St. Catharines, Ontario in 1986 with a Bachelor of Science in Geological<br>Sciences, and from Queen’s University in Kingston, Ontario in 1995 with a Master of Science in Environmental Engineering.
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4. I am a professional geologist and member in good standing with the Engineers and Geoscientists British<br>Columbia (No. 64426), the Association of Professional Geoscientists of Nova Scotia (No. 021), and the Association of Professional<br>Engineers, Geologists and Geophysicists of the Northwest Territories and Nunavut (No. 1624).
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5. I have practiced my profession for over 30 years. I have worked for mid- and large-size mining companies<br>where I have acted in senior technical and management roles, in senior environmental consulting roles, and providing technical advice<br>and subject matter expertise. My key areas of expertise include feasibility-level study reviews; NI 43-101 technical report writing and<br>review; environmental, social, and governance (ESG) due diligence for proposed mining operations and acquisitions, and the management<br>of environmental impact assessments (EIAs) and permitting applications to support the construction, operations, and closure of mining<br>projects. Recently, I acted as Qualified Person for environmental/sustainability sections in the following project reports: “Volcan<br>Project, NI 43-101 Technical Report on Preliminary Economic Assessment, Tierra Amarilla, Atacama Region, Chile”; “Santo Tomás<br>Copper Project, NI 43-101 Technical Report and Preliminary Economic Assessment, Northern Sinaloa State, Mexico”; "Tolillar<br>Project NI 43-101 Technical Report on Preliminary Economic Assessment, Salta Argentina"; “Santo Domingo Project NI 43-101 Technical<br>Report on Feasibility Study Update, Atacama Region, Chile”; and “Cerro Las Minitas Project NI 43-101 Technical Report Preliminary<br>Economic Assessment, Durango State, Mexico.
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6. I have read the definition of “Qualified Person” set out in the National Instrument 43-101<br>Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional<br>association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections<br>of the Technical Report that I am responsible for preparing.
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7. I have not visited the project site.
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8. I am responsible for Sections 1.18, 2.4.6, 3.2, 12.4, 20, 25.13, 25.17.2.9, 25.17.3.7, 26.7.3, 26.8, and<br>27 of this Technical Report.
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9. I am independent of New Pacific Metals Corp.as independence is defined in Section 1.5 of NI 43-101.
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10. I have not been previously involved with the Carangas Project.
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11. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared<br>in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, information and belief,<br>the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be<br>disclosed to make those sections of the Technical Report not misleading.
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Dated: August 13, 2026

“Signed and sealed”

James Millard, P.Geo.

CERTIFICATE OF QUALIFIED PERSON

Jinxing Ji, P.Eng.

I, Jinxing Ji, P.Eng., certify that:

1. I am employed as Consulting Metallurgist with JJ Metallurgical Services Inc. with an office at 7547Lambeth Drive, Burnaby, British Columbia, Canada, V5E 4A5.
2. This certificate applies to the technical report titled “Carangas Project: NI 43-101 TechnicalReport and Preliminary Economic Assessment, Oruro, Bolivia” that has an effective date July 16, 2026, and report date of August 13,2026 (the “Technical Report”).
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3. I graduated from Shanghai University in China with a bachelor’s in engineering Metallurgy in1982 and a master’s in engineering Metallurgy in 1985 and from The University of British Columbia in Canada with Ph.D. Metallurgyin 1993.
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4. I am a registered professional engineer (license# 59035) in good standing of the Engineers and GeoscientistsBC (EGBC) in the province of British Columbia, Canada with core competency in metallurgy.
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5. I have practiced my profession in the mining industry continuously since 1993. My relevant experiencesinclude mineral/metallurgical testing, process development, base-metal metallurgy, precious metal metallurgy, and process plant design,commissioning, optimization and operational support.
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6. I have read the definition of “Qualified Person” set out in the National Instrument 43-101Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professionalassociation and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sectionsof the Technical Report that I am responsible for preparing.
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7. I visited the Carangas Property in Bolivia from May 21 to 23, 2022.
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8. I am responsible for Sections 1.12, 2.4.4, 12.2, 13, 25.7, 25.17.2.4, 25.17.3.4, 26.3, and 27 of thisTechnical Report.
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9. I am independent of New Pacific Metals Corp. as independence is defined in Section 1.5 of NI 43-101.
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10. I have had prior involvement with the Carangas project as an independent metallurgical consultant forthe mineral/metallurgical testing under “Carangas Silver-Gold Project – Department of Oruro, Bolivia – NI 43-101 MineralResource Estimate Technical Report, New Pacific Metals Corp.” with affective date of 25 August 2023.
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11. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have beenprepared in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, informationand belief, the sections of the Technical Report for which I am responsible contain all scientific and technical information that is requiredto be disclosed to make those sections of the Technical Report not misleading.
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Dated: August 13, 2026

“Signed and sealed”

Jinxing Ji, P.Eng.

CERTIFICATEOF QUALIFIED PERSON****Anderson G. Candido, FAusIMM

I, Anderson G. Candido, FAusIMM, certify that:

1. I am employed as a Principal Resource Geologist with SLR Advisory Brasil Ltda (SLR), of Rua Antonio de<br>Albuquerque 330, 8th Floor, Belo Horizonte, MG, 30112-010, Brazil.
2. This certificate applies to the technical report titled “Carangas Project: NI 43-101 Technical Report<br>and Preliminary Economic Assessment, Oruro, Bolivia” that has an effective date July 16, 2026, and report date of August 13,<br>2026 (the “Technical Report”).
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3. I graduated from Ouro Preto Federal University with a Bachelor of Science in Geology in 2003.
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4. I am registered as a Professional Geologist in the Australasian Institute of Mining and Metallurgy (Fellow<br>Member No. 990424 “FAusIMM”).
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5. I have practiced my profession continuously for 22 years with experience in Geology and Mineral Resource<br>Evaluation, Exploration geological works, and Geology Development and Operations.
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6. I have read the definition of “Qualified Person” set out in the National Instrument 43-101<br>Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional<br>association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections<br>of the Technical Report that I am responsible for preparing.
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7. I visited the project site from March 27 to March 30, 2023.
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8. I am responsible for Sections 1.3–1.11, 1.13, 2.4.1, 4–11, 12.1, 14, 25.2–25.6, 25.8,<br>25.17.2.1–25.17.2.3, 25.17.2.5, 25.17.3.1–25.17.3.3, 25.17.3.5, 26.2, 26.4, and 27 of this Technical Report.
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9. I am independent of the New Pacific Metals Corp.as independence is defined in Section 1.5 of NI 43-101.
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10. I have been previously involved with the Carangas Project. I was the Geology QP for the 2024 PEA Technical<br>report.
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11. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared<br>in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, information and belief,<br>the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be<br>disclosed to make those sections of the Technical Report not misleading.
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Dated: August 13, 2026

“Signed and sealed”

Anderson G. Candido, FAusIMM

Page 1 of 1

CERTIFICATE OF QUALIFIED PERSON

Marc Schulte, P.Eng.

I, Marc Schulte, P.Eng., certify that:

1. I am employed as Vice President of Engineering and Operations with Moose Mountain Technical Services,<br>(MMTS), with an office address of #210 1510 2nd Street North Cranbrook, BC V1C 3L2.
2. This certificate applies to the technical report titled “Carangas Project: NI 43-101 Technical Report<br>and Preliminary Economic Assessment, Oruro, Bolivia” that has an effective date July 16, 2026, and report date of August 13,<br>2026 (the “Technical Report”).
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3. I graduated from the University of Alberta with a Bachelor of Science in Mining Engineering in 2002.
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4. I am a member of the self-regulated association Engineers and Geoscientists BC (No. 54035).
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5. I have worked as a mining engineer for over 22 years since my graduation from university. Throughout my<br>career I have worked on numerous open pit base and precious metals projects, within project engineering studies and within mining operations,<br>on mineral reserve estimates, mine planning, and mine cost estimates.
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6. I have read the definition of “Qualified Person” set out in the National Instrument 43-101<br>Standards of Disclosure for Mineral Projects (“NI 43-101”) and certify that by virtue of my education, affiliation to a professional<br>association and past relevant work experience, I fulfill the requirements to be a “Qualified Person” for those sections<br>of the Technical Report that I am responsible for preparing.
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7. I visited the project site on March 16 and March 17, 2026.
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8. I am responsible for Sections 1.14, 2.4.3, 15, 16, 21.2.3, 21.3.3, 25.9, 25.17.2.6, 26.5.1, and 27 of<br>this Technical Report.
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9. I am independent of New Pacific Metals Corp. as independence is defined in Section 1.5 of NI 43-101.
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10. I have been previously involved with the Carangas Project. I was the Mining QP for the 2024 PEA Technical<br>report.
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11. I have read NI 43-101 and the sections of the Technical Report for which I am responsible have been prepared<br>in compliance with that Instrument. As of the effective date of the Technical Report, to the best of my knowledge, information and belief,<br>the sections of the Technical Report for which I am responsible contain all scientific and technical information that is required to be<br>disclosed to make those sections of the Technical Report not misleading.
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Dated: August 13, 2026

“Signed and sealed”

Marc Schulte, P.Eng.

Page 1 of 1

Important Notice

This report was prepared as National Instrument 43-101 Technical Report for New Pacific Metals Corp. (NPM) by Ausenco Engineering Canada ULC, Ausenco Sustainability ULC (collectively, Ausenco), Moose Mountain Technical Services (MMTS), and SLR Advisory Brasil Ltda. (SLR), collectively the Report Authors. The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in the Report Authors’ services, based on i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by NPM subject to terms and conditions of its contracts with each of the Report Authors. Except for the purposes legislated under Canadian provincial and territorial securities law, any other uses of this report by any third party are at that party’s sole risk.

Carangas Project
NI 43-101 Technical Report and Preliminary Economic Assessment July 16, 2026
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Table of Contents

1 Summary 1
1.1 Introduction 1
1.2 Terms of Reference 1
1.3 Property Description and Location 1
1.4 Mineral Tenure, Surface Rights, Water Rights, Royalties and Agreements 2
1.5 Accessibility, Climate, Local Resources, Infrastructure and Physiography 2
1.6 History 3
1.7 Geology and Mineralization 3
1.8 Deposit Types 4
1.9 Exploration 4
1.10 Drilling 5
1.11 Sampling Preparation and Security 5
1.12 Mineral Processing and Metallurgical Testwork 6
1.13 Mineral Resource Estimate 7
1.14 Mining Methods 9
1.15 Recovery Methods 11
1.16 Project Infrastructure 15
1.16.1 On-Site and Off-Site Infrastructure 15
1.16.2 Tailings Storage Facility (TSF) 15
1.17 Market Studies and Contracts 15
1.18 Environmental, permitting and social considerations 16
1.18.1 Environmental Considerations 16
1.18.2 Closure and Reclamation Considerations 17
1.18.3 Permitting Considerations 17
1.18.4 Social Considerations 18
1.19 Capital and Operating Cost 18
1.19.1 Capital Cost Estimate 19
1.19.2 Operating Cost Estimate 20
1.20 Economic Analysis 21
1.20.1 Economic Summary 21
1.20.2 Sensitivity Analysis 22
1.21 Conclusions & Recommendations 22
Carangas Project Page i
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NI 43-101 Technical Report and Preliminary Economic Assessment July 16, 2026
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2 Introduction 23
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2.1 Introduction 23
2.2 Qualified Persons 23
2.3 Terms of Reference 24
2.4 Site Visits and Scope of Personal Inspection 24
2.4.1 Personal Inspection Details by Anderson G. Candido, FAusIMM 24
2.4.2 Personal Inspection Details by Scott Elfen, PE, P.Eng. 24
2.4.3 Personal Inspection Details by Marc Schulte, P.Eng. 24
2.4.4 Personal Inspection Details by Jinxing Ji, P.Eng. 25
2.4.5 Personal Inspection Details by Kevin Murray, P.Eng. 25
2.4.6 Personal Inspection Details by James Millard, P.Geo. 25
2.5 Sources of Information 25
2.5.1 Information and Data 25
2.5.2 Previous Technical Reports 25
2.6 Effective Dates 25
2.7 Currency, Units, Abbreviations and Definitions 26
3 Reliance on Other Experts 32
3.1 Introduction 32
3.2 Environmental, Permitting, Closure, and Social and Community Impacts 32
4 Property Description and Location 33
4.1 Property Description 33
4.2 Project Ownership 34
4.3 Mineral Tenure 35
4.4 Property Agreements 38
4.5 Project Risks and Uncertainties 38
5 Accessibility, Climate, Local Resources, Infrastructure and Physiography 39
5.1 Physiography and Climate 39
5.2 Accessibility 39
5.3 Local Resources and Infrastructure 41
6 History 42
6.1 Regional History 42
6.2 Property Exploration History 42
6.3 Historical Resource Estimates 43
6.4 Production 43
Carangas Project Page ii
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NI 43-101 Technical Report and Preliminary Economic Assessment July 16, 2026
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7 Geological Setting and Mineralization 44
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7.1 Regional Geology 44
7.2 Project Geology 46
7.3 Deposit Descriptions 49
7.3.1 Upper Silver Zone 50
7.3.2 Middle Zinc Zone 50
7.3.3 Lower Zone 50
8 Deposit Types 53
8.1 Deposit Model 54
8.2 Deposit Mineralization 54
9 Exploration 55
9.1 Sampling and Mapping 55
9.2 Geophysics 56
9.3 Exploration Potential 57
10 Drilling 58
10.1 Introduction 58
10.1.1 Drill Methods 61
10.1.2 Drilling Locational Data 61
10.1.3 Logging Procedures 61
10.1.4 Recovery 62
10.1.5 Sample Length/True Thickness 62
10.2 Comments on Drilling 62
11 Sample Preparation, Analyses, and Security 63
11.1 Introduction 63
11.2 Sample Collection 63
11.3 Assay Laboratory Sample Preparation and Analysis 65
11.4 Bulk Density 66
11.5 Quality Control Data 67
11.5.1 Certified Reference Materials (CRMs) 68
11.5.2 Blank Control Samples 69
11.5.3 Duplicate Samples 71
11.5.4 Umpire Laboratory Samples 74
11.6 Security and Storage 76
11.7 Comment on Sample Preparation, Analyses, Security and QA/QC 77
11.7.1 Recommendations 77
12 Data Verification 78
12.1 Geological and Mineral Resource Data Verification 78
12.1.1 Verification performed by the QP 78
Carangas Project Page iii
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NI 43-101 Technical Report and Preliminary Economic Assessment July 16, 2026
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12.2 Metallurgical Data Verification 81
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12.3 Geotechnical Data Verification 82
12.4 Socio-environmental data verification 82
13 Mineral Processing and Metallurgical Testing 83
13.1 Introduction 83
13.2 Historical Metallurgical Testwork (2022–2023) 84
13.2.1 Testwork of Flotation and Cyanide Leach (2022–2023) 84
13.2.2 Comminution Testwork (2023) 87
13.3 Recent Metallurgical Testwork (2024–2025) 88
13.3.1 Sample Selections and Head Assays 88
13.3.2 Mineralogy of the Oxidized Silver/Lead/Zinc Mineralized Samples 91
13.3.3 Bulk Flotation of the USZ Oxidized Composite Sample 93
13.3.4 Sequential Selective Flotation of the USZ Transitional Composite Sample 95
13.3.5 Sequential Selective Flotation of the USZ Sulfide Composite Sample 101
13.3.6 Sequential Selective Flotation of the USZ LOM Composite Sample 109
13.3.7 Sequential Selective Flotation of the Second USZ LOM2 Composite Sample 121
13.3.8 Metallurgical Testing for the Sample from the Lower Gold Zone 133
13.4 Metallurgical Variability 138
13.5 Deleterious Elements 140
13.6 Recovery Estimates 141
13.7 Comments on Mineral Processing and Metallurgical Testing 142
14 Mineral Resource Estimates 145
14.1 Introduction 145
14.2 Resource Database 145
14.2.1 Sample Data 146
14.2.2 Bulk Density Data 146
14.3 Depletion Areas 147
14.4 Geological Models 147
14.5 Resource Assays 149
14.6 Grade Capping/Outlier Restrictions 150
14.7 Compositing 152
14.8 Estimation/Interpolation Methods 153
14.8.1 Block Model Strategy and Analysis 153
14.8.2 Grade Interpolation 153
14.9 Density Assignment 155
14.10 Block Models 156
Carangas Project Page iv
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NI 43-101 Technical Report and Preliminary Economic Assessment July 16, 2026
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14.11 Reasonable Prospects for Eventual Economic Extraction 157
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14.11.1 Input Parameters for Resource Calculation 157
14.11.2 Pit Optimization Disclaimer 158
14.12 Classification of Mineral Resources 158
14.13 Block Model Validation 160
14.14 Mineral Resource Statement 164
14.15 Factors That May Affect the Mineral Resource Estimate 166
15 Mineral Reserve Estimates 167
16 Mining Methods 168
16.1 Introduction 168
16.2 Key Design Criteria 171
16.2.1 Net Smelter Prices, Net Smelter Return, and Cut-off Grade 171
16.2.2 Mining Loss and Dilution 174
16.2.3 Production Rate Considerations 175
16.3 Pit Optimization 176
16.3.1 Ultimate Pit Limits 177
16.4 Pit Designs 178
16.4.1 Open Pit Phases 178
16.4.2 In Pit Haul Roads 178
16.4.3 Open Pit Contents 179
16.4.4 Open Pit Designs 180
16.5 Low Grade and Oxide Stockpile Design 186
16.6 Waste Rock Storage Facility Design 186
16.7 Ex-Pit Haul Roads 187
16.8 Mine Production Schedule 187
16.8.1 End of Period Figures 191
16.9 Mine Operations 197
16.10 Risks 198
17 Recovery Methods 199
17.1 Overview 199
17.2 Process Flowsheet 200
17.3 Plant Design 203
17.3.1 Phase 1 Design (Years 1-5) 207
17.3.2 Phase 2 Design (Years 6-8) 211
17.3.3 Phase 3 Design (Years 9-16) 211
17.3.4 Counter-Current Decantation 212
17.3.5 Merrill Crowe Circuit 212
17.3.6 Refinery 213
17.3.7 Cyanide Detoxification and Tailings Management 214
Carangas Project Page v
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NI 43-101 Technical Report and Preliminary Economic Assessment July 16, 2026
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17.4 Energy, Water and Process Materials Requirements 214
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17.4.1 Process Materials 214
17.4.2 Water Requirements 215
17.4.3 Air Requirements 216
17.4.4 Oxygen Requirements 216
17.4.5 Power requirements 216
17.4.6 Personnel Requirements 216
18 Project Infrastructure 217
18.1 Introduction 217
18.2 Site Access 218
18.3 Built Infrastructure 219
18.3.1 On-site Roads 219
18.3.2 Process Plant Building 219
18.3.3 Accommodation 219
18.3.4 Support Buildings 219
18.4 Stockpiles 219
18.5 Tailings Storage Facility (TSF) 220
18.5.1 TSF Design and Construction 224
18.5.2 Stability Analysis 225
18.5.3 Geotechnical Instrumentation and Monitoring 225
18.5.4 TSF Closure 225
18.6 Waste Rock Storage Facility 225
18.7 Power and Electrical 226
18.8 Fuel 227
18.9 Water Supply and Management 227
18.9.1 Water Supply 227
18.9.2 Water Management 227
18.10 Hazard Considerations 231
18.10.1 Seismic Risk 231
18.10.2 Geohazards (Landslides, Rockfall, Debris Flow) 231
18.10.3 Cryospheric Hazards (Permafrost, Snow, Ice, Freeze–Thaw) 232
18.10.4 Flooding and River-Related Hazards 232
18.10.5 Weather-Driven Hazards (Winter Storms, Heavy Snow, Extreme Precipitation) 232
18.10.6 Avalanche Hazards 233
18.10.7 Climate-Driven Change 233
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19 Market Studies and Contracts 234
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19.1 Market Studies 234
19.2 Commodity Price Projections 235
19.3 Contracts 235
20 Environmental Studies, Permitting, and Social or Community Impact 236
20.1 Socio-Environmental Considerations – Baseline and Supporting Studies 236
20.1.1 Watershed Characteristics of the Project Area 237
20.1.2 Hydrology and Water Sources 241
20.1.3 Groundwater and Surface Water Quality 243
20.1.4 Biodiversity Assessment 248
20.1.5 Ecosystem Functions 251
20.1.6 Ambient Air Quality 252
20.1.7 Community Assessments and Studies 253
20.2 Environmental Management and Monitoring 257
20.2.1 Baseline Considerations and Derisking 257
20.2.2 Water Supply and Management 257
20.2.3 Environmental Management and Monitoring Plans 258
20.3 Permitting Considerations 259
20.3.1 Environmental Permits 259
20.3.2 Mining Permits 260
20.4 Social Considerations 260
20.4.1 Framework agreement with the Community 260
20.4.2 Prior Consultation Process 261
20.4.3 Collaborative Work with Community 261
20.5 Closure and Reclamation Planning 262
20.5.1 Closure and Reclamation Plans 262
20.5.2 Closure Cost Estimates 264
21 Capital and Operating Costs 265
21.1 Introduction 265
21.2 Capital Costs 265
21.2.1 Overview 265
21.2.2 Basis of Estimate 266
21.2.3 Mine Capital Costs 266
21.2.4 Process Capital Costs 268
21.2.5 Infrastructure Capital Costs 268
21.2.6 Indirect Capital Costs 270
21.2.7 Owner (Corporate) Capital Costs 270
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21.2.8 Sustaining Capital 271
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21.2.9 Contingency 272
21.2.10 Closure and Reclamation Planning 272
21.3 Operating Costs 272
21.3.1 Overview 272
21.3.2 Basis of Estimate 273
21.3.3 Mine Operating Costs 274
21.3.4 Process Operating Costs 275
21.3.5 General and Administrative Operating Costs 276
21.3.6 Owner (Corporate) Operating Costs 277
21.3.7 Closure and Reclamation Planning 278
22 Economic Analysis 279
22.1 Forward-Looking Information 279
22.2 Methodologies Used 280
22.3 Financial Model Parameters 280
22.3.1 Assumptions 280
22.3.2 Taxes 281
22.3.3 Working Capital 281
22.3.4 Closure Costs and Salvage Value 281
22.3.5 Royalties 281
22.3.6 Off-site Costs 282
22.4 Economic Analysis 282
22.5 Sensitivity Analysis 288
23 Adjacent Properties 291
24 Other Relevant Data and Information 292
25 Interpretation and Conclusions 293
25.1 Introduction 293
25.2 Mineral Tenure, Surface Rights, Water Rights, Royalties and Agreements 293
25.3 Geology and Mineralization 293
25.4 Exploration 293
25.5 Drilling 294
25.6 Sample Preparation, Analyses and Security 294
25.7 Metallurgical Testwork 295
25.7.1 Metallurgical Composite Samples 295
25.7.2 Comminution 296
25.7.3 Flotation of the Mineralized Materials in the Silver Zone 296
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25.7.4 Metallurgical Testing of the Mineralized Materials in the Gold Zone 296
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25.7.5 Process Flowsheet for the Mineralized Materials in the Silver Zone 296
25.7.6 Process Flowsheet for the Mineralized Materials in the Gold Zone 297
25.8 Mineral Resource Estimate 297
25.9 Mining Methods 298
25.10 Recovery Plan 298
25.11 Infrastructure 298
25.12 Markets and Contracts 299
25.13 Environmental, Permitting and Social Considerations 299
25.14 Capital Cost Estimate 301
25.15 Operating Cost Estimate 301
25.16 Economic Analysis 301
25.17 Risks and Opportunities 301
25.17.1 Introduction 301
25.17.2 Risks 302
25.17.3 Opportunities 306
26 Recommendations 310
26.1 Introduction 310
26.2 Exploration and Mineral Resource 310
26.2.1 Exploration and Drilling 310
26.2.2 Geology Study, Mapping, and Prospecting 311
26.2.3 Estimated Budget for Recommendations 311
26.3 Mineral Processing and Metallurgical Testing 312
26.4 Mineral Resource Estimate 314
26.5 Mining 314
26.5.1 Mine Engineering 314
26.5.2 Open Pit Geotechnical Program 315
26.6 Recovery Methods 316
26.7 Infrastructure 317
26.7.1 Infrastructure Geotechnical 317
26.7.2 TSF and WRSF Analyses and Studies 318
26.7.3 Hydrogeology 319
26.7.4 Hydrology 320
26.7.5 Geochemistry 320
26.8 Environmental Studies 321
27 References 323
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List of Tables

Table 1-1: Carangas Deposit - Conceptual Pit^*^ Constrained Mineral Resource as of March 31, 2026 8
Table 1-2: PEA Mine Plan Production Summary 9
Table 1-3: Capital Cost Summary 19
Table 1-4: Operating Cost Summary 20
Table 1-5: Recommended Work Programs 22
Table 2-1: Report Contributors 23
Table 2-2: Abbreviations and Acronyms 26
Table 2-3: Units of Measurement 30
Table 4-1: Mining Rights of the Carangas Property 35
Table 5-1: Weather of Carangas Region 39
Table 7-1 Summary of Carangas Mineralized Zones 51
Table 9-1 Summary of Exploration Programs at Carangas 55
Table 10-1: Carangas Drilling History 58
Table 10-2: Drilling Intercepts Examples 61
Table 11-1: QA/QC Sample Status 68
Table 11-2: CRMs of the Carangas Project 68
Table 11-3: Statistical Summary for Duplicate Samples July 2021 – April 2023 71
Table 11-4 Statistical Summary for Umpire Duplicates Samples 75
Table 12-1: Drill Core Intervals Viewed 79
Table 13-1: Summary of the Completed Metallurgical Testwork 84
Table 13-2: Drill holes and Intervals of Five Composite Samples in 2022 85
Table 13-3: Head Assays of Five Composite Samples in 2022 85
Table 13-4: Drill Holes and Core Intervals of Three Composite Samples for Comminution Testing 87
Table 13-5: Specific Gravity, Rod Mill Work Index, Ball Mill Work and Abrasion Index of Three Composite Samples 87
Table 13-6: Ten Groups of Samples from the Upper Silver Zone and Lower Gold Zone 90
Table 13-7: Head Assays of the Six Composite Samples in 2024 91
Table 13-8: Mineral Compositions of the USZ Oxidized and USZ Transitional Composite Samples 91
Table 13-9: Operating Conditions of Rougher Flotation Tests for the USZ Oxidized Composite Sample 93
Table 13-10: Results of Rougher Flotation Tests for the USZ Oxidized Sample 94
Table 13-11: Operating Conditions of Rougher Flotation Tests for the USZ Transitional Sample 96
Table 13-12: Results of Rougher Flotation Tests for the USZ Transitional Composite Sample 96
Table 13-13: Results of the Cleaner Flotation Test for the USZ Transitional Composite Sample 100
Table 13-14: Results of the Locked Cycle Test 84 for the USZ Transitional Composite Sample 100
Table 13-15: Operating Conditions of Rougher Flotation Tests for the USZ Sulfide Composite Sample 101
Table 13-16: Results of Rougher Flotation Tests for the USZ Fresh (Sulfide) Composite Sample 102
Table 13-17: Operating Conditions of the Cleaner Flotation Tests for the USZ Sulfide Composite Sample 105
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Table 13-18: Results of Cleaner Flotation Tests for the USZ Sulfide Composite Sample 106
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Table 13-19: Operating Conditions of Rougher Flotation Tests in the Silver/Lead Circuit for the USZ LOM Composite Sample 110
Table 13-20: Results of Rougher Flotation Tests in the Silver/Lead Circuit for the USZ LOM Composite Sample 110
Table 13-21: Operating Conditions of the Cleaner Flotation Tests for the USZ LOM Composite Sample 114
Table 13-22: Results of the Cleaner Flotation Tests for the USZ LOM Composite Sample 115
Table 13-23: Operating Conditions of the Locked Cycle Test 55 for the USZ LOM Composite Sample 120
Table 13-24: Results of the Locked Cycle Tests 55 and 57 for the USZ LOM Composite Sample 121
Table 13-25: Operating Conditions of the Rougher Flotation Tests for the USZ LOM2 Composite Sample 122
Table 13-26: Results of the Rougher Flotation Tests for the USZ LOM2 Composite Sample 122
Table 13-27: Operating Conditions of the Cleaner Flotation Tests in the Zinc Circuit for the USZ LOM2 Sample 125
Table 13-28: Results of the Cleaner Flotation Tests for the USZ LOM2 Sample 126
Table 13-29: Conditions of the Locked Cycle Tests for the USZ LOM2 Composite Sample 130
Table 13-30: Results of the Locked Cycle Tests for the USZ LOM2 Composite Sample 131
Table 13-31: Results of the E-GRG Gravity Concentration Test for the LGZ LOM Composite Sample 134
Table 13-32: Expected Gravity Gold Recovery from a Commercial Operation 135
Table 13-33: Operating Conditions of the Rougher Flotation Tests for the LGZ LOM Composite Sample 135
Table 13-34: Results of the Rougher Flotation Tests for the LGZ LOM Composite Sample 136
Table 13-35: Operating Conditions of the Selective Flotation Tests to Produce a Copper-Bearing Gold Concentrate 136
Table 13-36: Results of the Selective Flotation Tests to Produce a Copper-Bearing Gold Concentrate 137
Table 13-37: Conditions and Results of the Cyanide Leach of the LGZ LOM Composite Sample 138
Table 13-38: Conditions and Results of the Cyanide Leach of the Bulk Gold Flotation Concentrate 138
Table 13-39: Contents of Arsenic, Bismuth, Mercury, Antimony and Tin in the Domain Samples Calculated from the Assay Data of the Selected Intervals 141
Table 13-40: Metal Recoveries and Concentrate Grades in the Silver Zone 142
Table 14-1: Density Statistics Table 146
Table 14-2: Univariate Statistics of Grade Composites, by Domain 150
Table 14-3: Top Cut Values into all Domains 151
Table 14-4: Carangas Grade Estimation Search Parameters 154
Table 14-5: Density Estimation Parameters 155
Table 14-6: Carangas Block Model Definition Parameters 156
Table 14-7: Commodity Prices Used in Resource Calculation 157
Table 14-8: Metal payable percentages 158
Table 14-9: Composite vs. Block Model Grade Statistical Validation 161
Table 14-10: 3D Volumetric Model Comparison 161
Table 14-11: Carangas Deposit - Conceptual Pit* Constrained Mineral Resource as of March 31, 2026 165
Table 16-1: PEA Mine Plan Production Summary 168
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Table 16-2: Net Smelter Price and Recoveries for Mine Planning 172
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Table 16-3: Cut-off Grade 173
Table 16-4: Operating Cost Inputs for Pseudoflow Pit Shells 177
Table 16-5: Designed Open Pit Contents 179
Table 16-6: Mine Production Schedule 190
Table 17-1: Plant Feed Grades 200
Table 17-2: Plant Recoveries 200
Table 17-3: Concentrate Plant Process Design Criteria 203
Table 17-4: Gold Plant Process Design Criteria 205
Table 17-5: Major Equipment Sizing 206
Table 17-6: Reagents Consumption for Concentrate Plant 214
Table 17-7: Reagents Consumption for Gold Plant 215
Table 18-1: Channel Design Parameters 228
Table 18-2: Pond Sizing Parameters 229
Table 18-3: Monthly Water Balance: Precipitation, Evaporation and Pond Inflows 230
Table 18-4: Monthly Water Balance: Pond Outflows 230
Table 19-1: Off-take Term Assumptions 234
Table 20-1: Characteristics of the Carangas Sub-Basin 239
Table 20-2: Surface Water Evaluation Sites 240
Table 20-3: Flow Measurements 242
Table 20-4: Groundwater Evaluation Sites 243
Table 20-5: Groundwater Quality Test Results 243
Table 20-6: Surface Water Quality (upstream) 245
Table 20-7: Surface Water Quality (Middle) 246
Table 20-8: Surface Water Quality – Downstream and Control Points 247
Table 20-9: Number of Abiotic and Biotic Records of Flora 248
Table 20-10: Shrubland and Grassland Species 249
Table 20-11: Number of Abiotic and Biotic Records of Flora in Wetlands 250
Table 20-12: Wetland Species 250
Table 20-13: Lotic Ecosystem Evaluation Site List 252
Table 20-14: Population by Condition of Unmet Basic Needs 254
Table 20-15: Infrastructure of the Carangas Health Facility 256
Table 20-16: Environmental Licensing System Categorization 259
Table 21-1: Capital Cost Summary 265
Table 21-2: Mine Area Capital Cost Summary 267
Table 21-3: Process Plant Capital Cost Breakdown 268
Table 21-4: On-site Infrastructure Capital Cost Breakdown 269
Table 21-5: Off-site Infrastructure Capital Cost Breakdown 269
Table 21-6: Indirect Capital Cost Breakdown 270
Table 21-7: Mine Area Capital Cost Summary 271
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Table 21-8: Infrastructure Sustaining Capital Costs 271
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Table 21-9: Operating Cost Summary 272
Table 21-10: Process Operating Costs - Initial Phase - Concentrate Production 275
Table 21-11: Process Operating Costs - Expansion Phase - Concentrate Production 275
Table 21-12: Process Operating Costs - Expansion Phase – Doré Production 276
Table 21-13: General and Administrative Costs - Initial Phase 276
Table 21-14: General and Administrative Costs - Concentrator Expansion Phase 277
Table 21-15: General and Administrative Costs - Gold Expansion Phase 277
Table 22-1: Working Capital Assumptions 281
Table 22-2: Off-Take Term Assumptions 282
Table 22-3: Economic Analysis Summary Table 283
Table 22-4: Cashflow Statement on an Annualized Basis 285
Table 22-5: Pre-Tax Sensitivity 289
Table 22-6: Post-Tax Sensitivity 290
Table 26-1: Recommended Work Program 310
Table 26-2 Estimated Budget for Exploration and Mineral Resource Recommendations 311
Table 26-3: Summary of field investigations quantity 317
Table 26-4: Summary of Laboratory Tests 317

List of Figures

Figure 1-1: Mine Production Schedule Summary 11
Figure 1-2: Concentrate Process Plant Flowsheet 13
Figure 1-3: Gold Process Plant Flowsheet 14
Figure 4-1: Carangas Project General Location Plan 33
Figure 4-2: Location of Carangas Mineral Rights 37
Figure 5-1: Carangas Project General Location Plan 40
Figure 7-1: Regional Geology Plan 45
Figure 7-2: Project Geology Plan 47
Figure 7-3: Carangas Local Geology 48
Figure 7-4: Carangas Diatreme Structure Section View – Looking NW 49
Figure 7-5: Mineralized Zones by Metal Zoning- Oblique Section 52
Figure 8-1: Deposit Model 54
Figure 9-1: IP Chargeability Anomalies of the Carangas Area 57
Figure 10-1: Drillhole Collar Location Plan 59
Figure 10-2: Drillhole Cross-section 60
Figure 11-1: Drill Core Box Example – Drill Hole DCAr0171 64
Figure 11-2: Core Cutting and Sample Bag 65
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Figure 11-3: Specific Gravity Measurement 67
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Figure 11-4: Control Chart for CDN-ME-1501 (Ag) (July 2021 – November 2022) 69
Figure 11-5: Control Chart for Coarse Blank Samples 70
Figure 11-6: Control Chart for Pulp Blank CDN-GEO-1901 70
Figure 11-7 Precision Plot of Field (1/4 core) Duplicates for Silver Assays 72
Figure 11-8 Quantile-Quantile Plot of Field (1/4 core) Duplicates for Silver Assays 72
Figure 11-9 Coarse Duplicate Precision Scatterplot – Silver Assays 73
Figure 11-10 Precision Plot of Pulp Duplicates for Silver Assays 74
Figure 11-11 Umpire Pulp Duplicates Precision Scatterplot for Silver Assays 75
Figure 11-12 Secure Core Yard Storage 76
Figure 12-1: Drill Core Mineralization Intercept Examples 79
Figure 12-2: Examples of Drill Hole Collar Field Registration 80
Figure 13-1: Locations of the Selected Intervals for Metallurgical Samples 89
Figure 13-2: Deportment of Lead Minerals for the USZ Oxidized and USZ Transitional Composite Samples 92
Figure 13-3: Deportment of Zinc Minerals for the USZ Oxidized and USZ Transitional Composite Samples 92
Figure 13-4: Relationship between Silver Recovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ Oxidized Composite Sample 94
Figure 13-5: Silver/Lead Enrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ Oxidized Composite Sample 95
Figure 13-6: Silver Recovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ Transitional Sample 98
Figure 13-7: Silver/Lead Enrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ Transitional Sample 98
Figure 13-8: Silver/Zinc Enrichment Ratio of the Silver/Lead Concentrate for the USZ Transitional Sample 99
Figure 13-9: Silver Recovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ Sulfide Composite Sample 102
Figure 13-10: Silver/Lead Enrichment Ratio of the Silver/Lead Concentrate for the USZ Sulfide Composite Sample 103
Figure 13-11: Silver/Zinc Enrichment Ratio of the Silver/Lead Concentrate for the USZ Sulfide Composite Sample 103
Figure 13-12: Silver Recovery and Silver Content in the Silver/Lead Concentrate for the USZ Sulfide Sample 107
Figure 13-13: Lead Recovery and Lead Content in the Silver/Lead Concentrate for the USZ Sulfide Sample 107
Figure 13-14: Silver/Lead Enrichment Ratio of the Silver/Lead Concentrate for the USZ Sulfide Sample 108
Figure 13-15: Silver/Zinc Enrichment Ratio of the Silver/Lead Concentrate for the USZ Sulfide Sample 108
Figure 13-16: Zinc Recovery and Zinc Content in the Zinc Concentrate for the USZ Sulfide Sample 109
Figure 13-17: Silver Recovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ LOM Sample 112
Figure 13-18: Silver/Lead Enrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ LOM Sample 112
Figure 13-19: Silver/Zinc Enrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ LOM Sample 113
Figure 13-20: Silver Recovery and Silver/Lead Concentrate Mass Pull for the USZ LOM Composite Sample 117
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Figure 13-21: Silver Recovery and Silver Content in the Silver/Lead Concentrate for the USZ LOM Composite Sample 117
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Figure 13-22: Silver/Lead Enrichment Ratio of the Silver/Lead Concentrate for the USZ LOM Composite Sample 118
Figure 13-23: Lead Recovery and Lead Content in the Silver/Lead Concentrate for the USZ LOM Composite Sample 118
Figure 13-24: Silver/Zinc Enrichment Ratio of the Silver/Lead Concentrate for the USZ LOM Composite Sample 119
Figure 13-25: Silver Recovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ LOM2 Sample 123
Figure 13-26: Silver/Lead Enrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ LOM2 Sample 123
Figure 13-27: Silver/Zinc Enrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ LOM2 Sample 124
Figure 13-28: Stage Zinc Recovery and Zinc Content in the Zinc Rougher Concentrate for the USZ LOM2 Sample 124
Figure 13-29: Silver Recovery and Silver Content in the Silver/Lead Concentrate for the USZ LOM2 Sample 127
Figure 13-30: Silver/Lead Enrichment Ratio of the Silver/Lead Concentrate for the USZ LOM2 Composite Sample 127
Figure 13-31: Lead Recovery and Lead Content in the Silver/Lead Concentrate for the USZ LOM2 Composite Sample 128
Figure 13-32: Silver/Zinc Enrichment Ratio of the Silver/Lead Concentrate for the USZ LOM2 Composite Sample 128
Figure 13-33: Zinc Recovery and Zinc Content in the Zinc Concentrate for the USZ LOM2 Composite Sample 129
Figure 13-34: Silver Recovery in the Silver/Lead Concentrate between the Cleaner Tests and the Locked Cycle Tests for the USZ LOM2 Composite Sample 132
Figure 13-35: Lead Recovery in the Silver/Lead Concentrate between the Cleaner Tests and the Locked-Cycle Tests for the USZ LOM2 Composite Sample 132
Figure 13-36: Zinc Recovery in the Zinc Concentrate between the Cleaner Tests and the Locked Cycle Tests for the USZ LOM2 Composite Sample 133
Figure 13-37: Size-by-Size Gravity Recoverable Gold for the LGZ LOM Composite Sample 134
Figure 13-38: Silver Recovery and Silver Content in the Silver/Lead Concentrate for the Silver Zone 143
Figure 13-39: Lead Recovery and Lead Content in the Silver/Lead Concentrate for the Silver Zone 144
Figure 14-1: Three-Dimensional View of the Carangas Geological Model 148
Figure 14-2: Ag Log Histogram for 1.5 m Composites 149
Figure 14-2: Ag and Au Probability Plots with capping threshold 149
Figure 14-3: Length Histogram for Raw Assay Intervals 151
Figure 14-4: Estimation Density Histogram Validation 152
Figure 14-5: Classified Mineral Resources Block Model – Section DCAr0094 155
Figure 14-6: Swath Plot along X, Y, Z Direction for Ag (g/t) Validation 160
Figure 14-7: Silver (g/t) Grade Section View Validation of Block Model – Section 22 162
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Figure 14-8: Silver (g/t) Grade Section View Validation of Block Model – Section DCAr0094 163
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Figure 16-1: Mine Operations General Arrangements 169
Figure 16-2: Mining Loss and Dilution Application on 3800 masl bench 175
Figure 16-3: Pseudoflow Pit Shell Resource Contents by Case 178
Figure 16-4: Designed Open Pit Contents 180
Figure 16-5: Ultimate Pit Design, P621 182
Figure 16-6: Phased Pit Designs 183
Figure 16-7: Pit Designs, EW Section View, 7905350N 184
Figure 16-8: Pit Designs, NS Section View, 539150E 185
Figure 16-9: Annual Mill Feed Tonnes and Grade 188
Figure 16-10: Annual Material Mined and Waste Mining Ratio 189
Figure 16-11: Pit Phases Mined 189
Figure 16-12: End of Period Mine Production Schedule, Year -1 191
Figure 16-13: End of Period Mine Production Schedule, Year 1 192
Figure 16-14: End of Period Production Schedule, Year 3 193
Figure 16-15: End of Period Production Schedule, Year 5 194
Figure 16-16: End of Period Production Schedule, Year 9 195
Figure 16-17: End of Period Production Schedule, Year 19 196
Figure 17-1: Concentrate Process Plant Flowsheet 201
Figure 17-2: Gold Process Plant Flowsheet 202
Figure 18-1: General Arrangement 218
Figure 18-2: TSF General Arrangement 220
Figure 18-3: TSF Starter Facility 222
Figure 18-4: TSF Ultimate Facility 223
Figure 18-5: Cross-section for CDSF Embankment 224
Figure 18-6: Ultimate WRSF Design 226
Figure 20-1: Carangas Sub-Basin Hydrographic Map 238
Figure 20-2: Surface Water Evaluation Sites 241
Figure 22-1: Sensitivity Analysis 288
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1 Summary
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1.1 Introduction
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Ausenco Engineering Canada ULC, and Ausenco sustainability (collectively, Ausenco) has prepared a preliminary economic assessment (PEA) and associated technical report for New Pacific Metals Corp. (NPM) on the Carangas Project, located in the Carangas District, situated on the western part of the Oruro Department, in Bolivia. The technical report was prepared in accordance with the Canadian disclosure requirements of National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101) and Form 43-101 Form 1 – Technical Report.

The responsibilities of the engineering consultants are as follows:

· Ausenco was responsible for the process plant, infrastructure, tailings, waste rock, water management,<br>environmental, permitting, and community, cost estimation, financial modelling, and overall report compilation.
· Moose Mountain Technical Services (MMTS) was responsible for mining, stockpiles, haul roads, and mining-related<br>cost estimation.
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· SLR Advisory Brasil Ltda (SLR) was responsible for the property, geology, exploration, drilling, and sampling<br>in addition to the mineral resource estimation.
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· JJ Metallurgical Services Inc. was responsible for the metallurgical testwork and recoveries.
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1.2 Terms of Reference
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All measurement units used in this report are SI/metric units unless otherwise noted. Currency is expressed in United States dollars (US$ or USD) unless otherwise noted.

Mineral resources are reported in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and Mineral Reserves (CIM, 2014) and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (CIM, 2019).

1.3 Property Description and Location

The Carangas Property is located in the Carangas district, situated in Bolivia's western part of the Oruro Department, approximately 190 kilometers (km) southwest of Oruro City. The property is currently held by Minera Granville SRL. (Granville), a private Bolivian company, and comprises two Prospecting and Exploration Licenses (PELs), namely Granville and Colapso, covering a total area of 39 km². There is currently a camp for geologists to support exploration activities. Access is via the existing 36 km five-meter wide access road from the paved main Highway 12 to the camp.

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1.4 Mineral Tenure, Surface Rights, Water Rights, Royalties and Agreements
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The Property consists of two PELs granted to Minera Granville S.R.L. (Granville) by the Bolivian Mining Administrative Jurisdictional Authority (AJAM). Granville is a private Bolivian company 100% owned by Bolivian nationals. Each PEL has a five-year validity term, with provisions for a single three-year extension.

The Carangas Project is located within 50 km of the Bolivian border with Chile. In line with many South American countries, Bolivia does not permit foreign entities to own property within 50 km of international borders. Property owners in the Frontier Area are, however, permitted to enter into mining association agreements with third parties, including foreign entities, for the development of mining activities under Bolivian Law No. 535 on Mining and Metallurgy. In April 2021, New Pacific Metals Corp. (NPM) entered into a Mining Association Contract (MAC) with Granville to jointly explore and develop the Property under applicable Bolivian laws and pursuant to the terms and conditions of the MAC, whereby NPM acquire 98% profits generated from the Project. NPM will cover all costs related to the exploration, development, and mining of the Project, with the majority of the profits from mining production going to NPM and a smaller portion allocated to Granville. As the holder of the mineral title to the Property, Granville will be responsible for permitting matters to ensure the Property remains in good standing under applicable Bolivian laws. The agreement has a term of 30 years and is renewable for an additional 15 years.

Before April 2021, NPM had been working in the area, completing surface and underground (UG) mapping and recognizing the main geological structures and mineralization settings.

The Qualified Person (QP) is not aware of any environmental liabilities on the Property. NPM has all the required permits to conduct the proposed work on the Property. The QP is not aware of any other significant factors and risks that may affect access, title, or the right or ability to perform the proposed work program on the Property.

1.5 Accessibility, Climate, Local Resources, Infrastructure and Physiography

The Carangas Project is reached by travelling from Oruro along a paved highway to Sabaya and then following a 6 km paved road followed by a 30 km gravel road to site. This route allows trucks, equipment, and personnel to reach the site year-round without major issues. Oruro, the closest major city, provides skilled workers and supplies, while the nearby Carangas community is small but lies directly next to the Project area.

The Project sits high on the Andean plateau at approximately 3,900 to 4,100 meters above sea level (masl). The landscape is mostly gentle and open, marked by two large volcanic domes separated by a valley. Vegetation is sparse, typical of the dry, cool climate at high altitude. Rainfall mainly occurs during a short wet season, but overall weather conditions allow fieldwork and mining activities to continue year-round.

Local resources can support both current exploration and potential future development. Water is assumed to be available from water wells to be drilled upstream of the pit, and electrical power is available from the proposed 230 kV transmission line connecting the Project site with the Pagador substation. The site already has an exploration camp with accommodation, core logging facilities, storage, and reliable communications. Although NPM holds exploration rights, the land surface is owned by local communities, and any future expansion, such as building the plant or tailings facilities, will require new agreements and continued collaboration with those communities.

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1.6 History
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Mining activities in the Carangas district began in the late 16th century in the Spanish colonial era. During that time, mining activities were mainly focused on oxide materials and native silver. Currently, widespread ruins of historical mine workings are visible in the East Dome and West Dome, historically known as San Antonio and Espiritu Santo hills.

Following the decline of the Spanish colonial era, mining activities in the Carangas area diminished. In the 20th century, ownership of the Property was transferred between various international and local Bolivian mining companies. Notably, in the early 20th century, mining operations were revived by Moritz Hochschild and Federico Alhfeld, a German geologist regarded as the father of Bolivian geology, who was working on the Property in 1923.

There has been very limited historical mineral exploration at the Property. Compañía Minera del Sur S.A. (COMSUR), a local Bolivian mining company, conducted the earliest recorded exploration, including channel sampling in the underground workings of the San Jose, Orcko Tunku, and San Antonio adits in 1985. It collected over 350 samples with an average grade of 64 grams per tonne (g/t) of silver (Ag). Llicancabur Mining Ltda. (Llicancabur), a local Bolivian mining company completed a total of 1,001 meters (m) in nine reverse circulation holes in 1995, and COMSUR drilled 914.2 millimeters (mm) in 6 diamond drill holes (DDH) in 2000 (Lopez-Montaño, 2019).

The Carangas Project published its first Mineral Resource Estimate (MRE) in 2023, which served as the foundation for the 2024 PEA Technical Report. Since that time, no additional drilling or geological updates have been completed within the project area. For the current PEA update, only the silver equivalent (AgEq) and net smelter return (NSR) values were revised to reflect updated metal prices and related technical assumptions. A new resource constraint pit shell was generated using these updated inputs, and this pit now forms the basis of the 2026 Mineral Resource statement. All previously disclosed Mineral Resource estimates are therefore superseded by the estimates presented in this report.

1.7 Geology and Mineralization

The Property sits in the South American Epithermal-Porphyry Belt, which features a geological sequence that includes Jurassic granites and the volcanic rocks of the Negrillos and Carangas Formations of Tertiary age. The Negrillos Formation consists of eroded lavas, tuffs, and volcanic breccias from ancient volcanic cones. Above the Negrillos Formation, the Carangas Formation includes rhyolitic to rhyo-dacitic intrusive dykes, lithic tuffs, phreatomagmatic breccias intercalated with fluvial sediments in the upper portion, and andesitic volcanoclastic rocks in the lower portion.

The Carangas area is interpreted as a grand volcanic caldera system of the Tertiary age. The Property is located at the southwest corner of the Carangas basin. It geomorphologically comprises two prominent hills, the West Dome and the East Dome, with a fluvial valley between them called the Central Valley. In addition, there is a small hill known as South Dome near the south end of the Central Valley. At the Property's surface, silver-lead-zinc mineralized vein structures predominantly strike in a west-northwest direction with steep dips, either sub-vertically or slightly dipping to the south or the north. In addition, some vein sets trend in northerly and northeast directions. To a depth below the shallow silver-lead-zinc horizon, mineralization is dominated by gold plus a minor amount of silver and copper in the lower portion of the mineralized system.

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Based on drilling data, the West Dome and Central Valley areas are interpreted as a diatreme structure with an inverted-cone shape, filled with breccias of phreatomagmatic origin and rhyodacitic intrusive dykes. At the top of West Dome, unlithified sandy sediments with horizontal beddings intercalated with phreatomagmatic breccias of altered rhyolitic and older volcanoclastic clasts are well exposed on the surface, indicating a volcanic maar environment. The intrusion of magma, once reaching the meteoric water level near the surface, led to a series of intense explosive eruptions and fracturing, which in turn generated abundant open spaces, including cracks and pores in breccias, favourable for the circulation of hydrothermal fluids and the deposition of sulfide minerals of metals. Three zones of mineralization can be recognized, each with different metals. The Upper Silver Zone is near-surface and is dominated by silver, with moderate amounts of lead and zinc. Below the upper zone, the Middle Zinc Zone is dominated by zinc plus minor silver and lead. The Lower Gold Zone is dominated by gold, with small amounts of silver, copper, and zinc.

1.8 Deposit Types

The Carangas Project is interpreted to be part of a large epithermal precious-metal system developed within the Central Andean volcanic arc of western Bolivia, a region well known for hosting silver- and gold-rich epithermal deposits. Mineralization at Carangas is associated with tertiary volcanic rocks, including volcanic domes, breccias, and related intrusive phases, and shows a clear vertical metal zoning with silver-dominant mineralization near surface and gold-dominant mineralization at depth. This vertical zonation, together with widespread hydrothermal brecciation, strong alteration, and a close spatial relationship to volcanic centers, is consistent with a low- to intermediate-sulfidation epithermal deposit model. Regional studies of the Andean Altiplano indicate that such systems commonly form above deep magmatic heat sources, with mineralization focused along structural corridors and caldera-related features. At Carangas, geophysical data, drilling, and surface exposure suggest a vertically extensive hydrothermal system that remains open laterally and at depth, consistent with epithermal systems elsewhere in the Central Andes (e.g., USGS Andean epithermal models; Sillitoe, 2010).

1.9 Exploration

The Carangas Project underwent a systematic exploration process, beginning with Granville's reconnaissance mapping and sampling in 2019. This initial phase was followed by detailed surface-underground mapping and sampling throughout 2020-2021. Exploration activities continued intermittently in 2022 and concluded with the sampling and mapping of previously inaccessible historical underground workings.

In 2020, Granville collected 383 rock chip samples from 55 outcrops. The samples were taken at two-meter intervals, approximately perpendicular to the strike direction of mineralization, covering a total length of 769 m. Out of these samples, 117 returned grades ranging from 30 to 2,350 g/t Ag, with an average grade of 160 g/t Ag. These samples were used as a guideline for further exploration programs.

The Property features historical underground mining workings. The NPM conducted surveys of all safe and accessible tunnels totalling, 2.4 km, all developed within the Carangas Formation. To date, a total of 425 samples have been collected. Among these samples, 112 (26.35%) returned assay results ranging from 30 to 1,060 g/t Ag, with an average grade of 122 g/t Ag.

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Furthermore, the NPM implemented systematic geophysical surveying programs, including a ground magnetometry survey and an Offset (3D) Bipole-Dipole IP-MT survey, from 2021 to 2023. The known mineralization system responds well to magnetic lows and IP chargeability highs, and multiple additional anomalies were identified.

1.10 Drilling

The NPM started exploration drilling in June 2021 and completed resource definition drilling at the end of April 2023. During that period, as many as five rigs were running at Carangas, and a total of 81,145 m were drilled in 189 holes. Maldonado Exploraciones, a contracted drilling company based in La Paz, Bolivia, conducted all drilling, divided into roughly four stages.

· Phase I: Drilling started on June 21, 2021, and concluded on September 24, 2021. Thirteen holes<br>were completed, totalling 3,790.4 m, to verify historical drill results and to test the lateral and depth extent of the known mineralization<br>exposed on the surface at West Dome and East Dome.
· Phase II: Drilling commenced on October 6, 2021, and was completed on December 17, 2021. In<br>this phase, 22 holes were drilled for a total of 9,420 m with the objective of testing mineralization covered by young sediments<br>in the Central Valley area.
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· Phase III: A resource definition drill program, was started on February 3, 2022, and completed on<br>December 14, 2022. Five drill rigs were employed to rapidly define the mineral resource potential at Carangas. During this period,<br>a total of 50,311 m were drilled in 115 holes on a drill grid of approximately 50 m spacing, and most holes intersected broad mineralization.
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· Phase IV: Drilling progressed as a continuation of the 2022 resource definition drill program to infill<br>areas drilled in 2021-2022, as well as step out beyond these previously drilled areas. As of the end of April 2023, a total of 39<br>holes were completed for a total of 17,623.5 m during this phase of drilling.
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1.11 Sampling Preparation and Security
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NPM has established a series of working procedures and protocols regarding core logging, sampling, core quality assurance/quality control (QA/QC), and data validation, which include the regular submission of check samples to an umpire laboratory in Lima, Peru.

All drill holes were geologically logged and sampled by NPM field personnel at the company’s facilities in Carangas. Geological logging included detailed recording of lithology, alteration, mineralization, structure, and rock quality designation (RQD) measurements. Drill cores are stored in a secure core storage area at the Carangas camp for future checks and audits.

NPM personnel oversee the delivery of drill core and rock chip samples from the Carangas camp to the ALS Laboratories (ALS) in Oruro, Bolivia, for sample preparation. The pulp samples were then shipped to ALS in Lima, Peru, for geochemical analysis. ALS Oruro and ALS Lima are part of ALS Global, a commercial laboratory specializing in analytical geochemistry services, all of which are accredited to International Organization for Standardization (ISO)/ International Electrotechnical Commission (IEC) 17025:2017 and are independent of NPM.

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All drill core, rock chip, and grab samples are prepared using the following procedures: (1) crush to 70% less than 2 mm; (2) riffle split of 250 g; and (3) pulverize the split to more than 85% passing a 75 µm sieve.

NPM has established comprehensive QA/QC procedures covering every step of sampling, preparation, and geochemical analysis, including inserting certified reference materials (CRMs), blanks, and duplicates into regular sample sequences. The use of a reasonable number of different control samples is robust and the umpire laboratory check analysis provides a high level of reproducibility of the database.

The insertion ratio of control samples is 24%, which is higher than the industry benchmark (15% to 20%).

In the QP's opinion, the data acquisition, analysis, and validation comply with the best industry practices and are trustworthy for Mineral Resource estimates and technical reporting.

1.12 Mineral Processing and Metallurgical Testwork

Following completion of the first metallurgical testwork program in May 2023 with five mineralized samples, another six mineralized samples were collected in December 2023 from the upper silver zone and lower gold zone for the second metallurgical testwork program to support the preliminary economic assessment in 2024. In the upper silver zone, five mineralized samples were prepared using a large number of intervals from multiple drill holes.

· The first composite sample was nearly fully oxidized mineralization which contained 63 g/t silver (Ag),<br>0.46% lead (Pb) and 0.09% zinc (Zn). The open-circuit bulk flotation with two-stage upgrade generated a silver/lead concentrate which<br>contained 8,330 g/t silver and 27.5% lead with 74.8% silver recovery and 33.7% lead recovery.
· The second composite sample was partially oxidized mineralization which contained 55 g/t Ag, 0.45% Pb<br>and 0.66% Zn. The locked cycle selective flotation with two-stage upgrade produced a silver/lead concentrate containing 3,070 g/t<br>Ag and 19.4% Pb with 83.8% Ag recovery and 65.4% Pb recovery, and a zinc concentrate containing 208 g/t Ag and 45.1% Zn with 1.2% Ag recovery<br>and 21.1% Zn recovery.
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· The third composite sample was fresh (sulfide) mineralization which contained 55 g/t Ag, 0.39% Pb and<br>0.77% Zn. The open-circuit selective flotation with two-stage upgrade resulted in a silver/lead concentrate containing 2,130-4,880 g/t<br>Ag and 14.7-33.5% Pb with 80.1-87.3% Ag recovery and 83.1-87.5% Pb recovery, and a zinc concentrate containing 266 g/t Ag and 50.2% Zn<br>with 4.7% Ag recovery and 65.1% Zn recovery.
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· The fourth composite sample was a life-of-mine (LOM) composite, which was prepared at a ratio of 12.5%<br>fully oxidized sample, 2.5% partially oxidized sample and 85.0% fresh (sulfide) sample, and contained 52 g/t Ag, 0.38% Pb and 0.67% Zn.<br>The locked cycle selective flotation with two-stage upgrade generated a silver/lead concentrate containing 3,675 g/t Ag and 24.4% Pb with<br>82.1% Ag recovery and 74.1% Pb recovery, and a zinc concentrate containing 332 g/t Ag and 45.3% Zn with 6.8% Ag recovery and 69.3% Zn<br>recovery.
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· The fifth composite sample was a composite representing the first eight years of operation, which was<br>prepared at a ratio of 17.0% fully oxidized sample, 3.0% partially oxidized sample and 80.0% fresh (sulfide) sample, and contained 73<br>g/t Ag, 0.45% Pb and 0.73% Zn. The locked cycle selective flotation with two-stage upgrade produced a silver/lead concentrate containing<br>6,088 g/t Ag and 31.2% Pb with 81.2% Ag recovery and 67.7% Pb recovery, and a zinc concentrate containing 343 g/t Ag and 46.3% Zn with<br>4.5% Ag recovery and 61.2% Zn recovery.
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In the lower gold zone, one composite sample was prepared using a large number of intervals from multiple drill holes. This composite sample contained 1.03 g/t gold (Au) and 9.2 g/t Ag.

· A three-stage gravity concentration test recovered 67.3% Au. The gold recovery from a commercial gravity<br>concentration circuit is expected to be 42-46%.
· The whole-ore cyanide leach resulted in 94.0% Au recovery and 50.9% Ag recovery.
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· The bulk flotation achieved 98.0% Ag recovery and 94.7% Ag recovery at 10.9% mass pull.
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· The bulk flotation concentrate was amenable to cyanide leach with average 95.3% Au recovery and 63.8%<br>Ag recovery.
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Three samples from the upper silver zone, lower silver zone and lower gold zone were subjected to the comminution testing. The measured rod mill work index, ball mill work index and abrasion index values were 10.1–12.3 kWh/t, 10.7–12.8 kWh/t and 0.038–0.075 g. These values indicate that these three samples were moderately hard and mildly abrasive.

1.13 Mineral Resource Estimate

The Mineral Resource of the Carangas Project was independently estimated based on the data provided by NPM as of June 1, 2023. The Mineral Resource estimate and underlying data comply with the guidelines of the CIM Definition Standards for Mineral Resources and Mineral Reserves (CIM, 2014) under NI 43-101. The QP considers the Mineral Resource estimate suitable for public reporting. The QP, Mr. Anderson Goncalves Candido, completed the Mineral Resources estimate.

The Mineral Resource estimate has an effective date of March 31, 2026 and was reported using a cut-off grade of 30 g/t of silver equivalent (AgEq) and is constrained within a conceptual open pit shell, assuming extraction will be conducted using an open pit mining method. The cut-off value was determined using technical factors including assumed costs, metallurgical recoveries, and metal prices considered reasonable by the QP.

Mineral Resources are individually reported for the three mineralization zones identified at the Project: the Upper Silver Zone, the Middle Zinc Zone, and the Lower Gold Zone. The results of the Mineral Resource estimate for the Carangas deposit are presented in Table 1-1. Mineral Resources that are not mineral reserves do not have demonstrated economic viability.

Below the conceptual pit constraint, gold-dominated mineralized material of similar size and grade to the reported Mineral Resources of the Lower Gold Zone exists. Gold mineralization remains open to the north and northeast at depth. The material below the Lower Gold Zone is not included in the Mineral Resource estimate.

The reported Mineral Resources have reasonable prospects for eventual economic extraction (RPEEE) using open-pit mining methods.

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Table1-1:      Carangas Deposit - Conceptual Pit^*^ Constrained Mineral Resource as of March 31,2026

Domain Category Density Tonnage AgEq Ag Au Pb Zn Cu
g/cm^3^ Mt Grade (g/t) Contained (Moz) Grade (g/t) Contained (Moz) Grade (g/t) Contained (Koz) Grade (%) Contained (Mlb) Grade (%) Contained (Mlb) Grade(%) Contained(Mlb)
Upper Silver Zone Indicated 2.09 121.6 70 272.4 44 173.9 0.06 220.0 0.34 923.3 0.65 1,729.6 0.01 34.91
Inferred 2.09 33.6 67 72.9 42 45.9 0.11 119.6 0.29 211.6 0.48 357.6 0.01 9.70
Middle Zinc Zone Indicated 2.29 38.7 41 51.3 12 15.1 0.06 68.8 0.37 314.1 0.81 688.0 0.02 12.96
Inferred 2.33 9.1 39 11.5 9 2.6 0.05 15.3 0.38 77.1 0.85 172.2 0.01 2.57
Lower Gold Zone Indicated 2.27 78.4 82 205.9 10 24.4 0.76 1,911.7 0.12 215.0 0.23 403.8 0.06 98.08
Inferred 2.33 11.0 80 28.3 10 3.6 0.70 248.5 0.15 36.2 0.34 82.4 0.05 11.78
Total Indicated 2.18 238.8 69 529.6 28 213.4 0.29 2,200.5 0.28 1,452.4 0.54 2,821.3 0.03 145.96
Inferred 2.18 53.8 65 112.7 30 52.1 0.22 383.4 0.27 324.9 0.52 612.3 0.02 24.05

* Notes:

1. CIM (2014) definitions were followed for Mineral Resources.
2. The QP for the Mineral Resource estimate, as defined by NI 43-101, is Anderson Candido, FAusIMM, Principal<br>Geologist at SLR. The QP is independent of NPM and the Carangas Project.
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3. Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.
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4. The results are presented undiluted and on a dry in-situ basis and are considered to have reasonable prospects<br>of economic viability.
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5. Mineral Resources are estimated at a cut-off grade of 30 g/t AgEq.
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6. AgEq formula is: AgEq g/t = Ag g/t + Au g/t * 80.49 + (Pb %*2204.6 /100 + Zn %*2866 /100 + Cu %* 8818.5<br>/100) / 1.318.
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7. Density varies from 2.09 in the upper zone to 2.33 in the lower zone.
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8. Mineral Resources are constrained by an optimized pit shell at a metal price of $41.00/oz Ag, $3,300.00/oz<br>Au, $1.00/lb Pb, $1.30/lb Zn, $4.00/lb Cu. Recovery of 81.6% Ag, 93.4% Au, 73.4% Pb, 66.9% Zn, 38.7% Cu.
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9. 1.85% of Mineral Resources are within the constrained pit but not within the mining concessions held by<br>NPM.
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10. The drilling database was closed on June 1, 2023.
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11. The numbers may not add due to rounding.
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1.14 Mining Methods
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The deposit is amenable to open-pit mining practices. Open pit mine designs, mine production schedules and mine capital and operating costs have been developed for the Carangas deposit at a PEA level of engineering. The Mineral Resources described in Section 14 form the basis of the mine planning, including Indicated and Inferred class resources.

Mine planning is based on conventional drill, blast, load, and haul open pit mining methods suited for the project location and local site requirements. The open pit activities are designed for two years of construction followed by sixteen years of mine operations and three years of post-pit mining stockpile rehandle to the mill. The subset of Mineral Resources contained within the designed open pits are summarized in Table 1-2, with a $13.50/t NSR cut-off and form the basis of the mine plan and production schedule.

Table 1-2:      PEAMine Plan Production Summary

Factor Value
PEA Silver Zone Mill Feed 200.1 Mt
Silver Zone Mill Feed NSR Grade $42.2/t
Silver Zone Mill Feed Ag Grade 36.2 g/t
Silver Zone Mill Feed Pb Grade 0.34 %
Silver Zone Mill Feed Zn Grade 0.64 %
PEA Gold Mill Feed 51.4 Mt
Gold Zone Mill Feed NSR Grade $75.7/t
Gold Zone Mill Feed Ag Grade 11.9 g/t
Gold Zone Mill Feed Au Grade 0.74 g/t
Waste Rock 363.4 Mt
Waste: Resource Ratio 1.4

Notes:

1. The PEA Mine Plan and Mill Feed estimates are a subset of the March 31, 2026, Mineral Resource estimates<br>and are based on open pit mine engineering and technical information developed at a Scoping level for the Carangas deposit.
2. PEA Mine Plan and Mill Feed estimates are mined tonnes and grade; the reference point is the primary crusher.<br>Mill Feed tonnages and grades include open pit mining method modifying factors, such as dilution and recovery.
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3. Net Smelter Prices (NSP) and metallurgical recoveries define the cut-off grade. NSPs include market price<br>assumptions of $40.0/oz Ag, $1,984/t Pb, $2,646/t Zn, $3,200/oz Au. Various smelter and refining terms, offsite costs, and a 5% royalty<br>(6% for Ag) derive NSPs of $17.5/oz Ag in Zn concentrate, $34.4/oz Ag in Pb concentrate, $34.8/oz Ag in doré, $3,028/oz Au in doré,<br>$1,070/t Pb, and $1,270/t Zn. Metallurgical recoveries of 6% Ag in Zn concentrate, 81.6% Ag in Pb concentrate, 60.4% Ag in doré,<br>93.4% Au in doré, 2.3% Pb in Zn concentrate, 73.4% Pb in Pb concentrate, 66.9% Zn in Zn concentrate, and 21.9% Zn in Pb concentrate<br>are applied.
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4. The chosen cut-off grade covers total operating costs of $13.50/t, which exceeds estimated PEA processing<br>and G&A cost estimates.
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5. Estimates have been rounded and may result in summation differences.
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Economic pit limits are determined using the Pseudoflow implementation of the Lerchs-Grossman algorithm. Selected pit limits are split up into five phases or pushbacks to target higher economic margin material earlier in the mine life. Upper benches of the open pit will be accessed via internal cut ramps on topography or via ramps left behind on phased pit walls. In-pit ramps will access material below the pit rim.

Pit designs are configured on 10 m bench heights, with a minimum of 8 m wide berms placed every two benches or double benching. Since no geotechnical test work or analysis has been completed on the bedrock, the applied bench face and inter-ramp angles, 67.5 degrees and 50 degrees, respectively, are scoping level assumptions based on the rock type and overall depth of the open pit.

The pit design requires mining of waste on Mining Concessions in the southern portion of the planned open pit that do not belong to the Company. These concessions include approximately 1.8% of the mineral resources that have been included in the economic analysis for this PEA. These Concessions are held by the state of Bolivia and are not currently available for tenure. Although the Company is actively working with the Bolivian government to obtain them, failure to do so, or to enter into a mining agreement on them could cause the Company to reevaluate the pit design and the outcome of this PEA.

Resources from the open pit will report to a Run-of-Mine (ROM) pad and primary crusher 0.5 km northeast of the pit rim. The mill will be fed with Silver Zone mineralized material from the pits at an average rate of 8.0 Mt/a (22 kt/d), increasing to 16.0 Mt/a (44 kt/d) in Year 6 of the Project. Between Years 9 and 16 of the Project, half the mill feed will be dedicated processing the gold zone mineralization.

Oxide resources will be stored in a stockpile 0.5 km east of the pit rim and rehandled to the crusher over the LOM, blended with non-oxide mill feed. Non-oxide resources, mined in excess of mill feed targets, will be stored in a low-grade stockpile 1.5 km southeast of the ROM pad and process plant and east of the open pit. This stockpile is planned to be completely reclaimed to the mill at the end of the mine life.

Waste rock will be placed in a side hill waste rock storage facility (WRSF) 0 to 2.5 km north of the open pit, or used in the construction of haul roads and the dam portion of the tailings facility, which sits 3.0 km north of the open pit.

The waste rock from the open pit has not been tested or analyzed for potential acid generation (PAG). It is assumed that PAG quantities will be small enough to be blended with larger quantities of non-acid-generating (NAG) waste rock for surface storage within the WRSFs.

Topsoil and overburden encountered at the top of the pits will be placed in a dedicated stockpile directly north of the open pit and kept salvageable for closure at the end of the mine life. These quantities have not been measured and the storage facility has not been designed for the PEA mine plan.

The mine production schedule is summarized in Figure 1-1 below.

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Figure 1-1:      MineProduction Schedule Summary

Source: MMTS, 2026

Contractor mining operations are planned, utilizing a diesel-powered mining fleet. Cost estimates for mining are based on a contractor quotation for this project, utilizing down-the-hole (DTH) drills for drilling, 0.20 kg/t target powder factor ANFO-based blasting, 9 m^3^ bucket size diesel hydraulic excavators for loading, and 90 t payload rigid-frame haul trucks for hauling, plus ancillary and service equipment to support the mining operations, including haul road and stockpile maintenance.

In-pit dewatering systems will be established for the pit. All surface water and precipitation in the open pit will be gravity drained or directed via pumps to ex-pit settling ponds directly outside the pit limits, where it will report to the broader project water management system.

Contractor cost estimates include investment in the mining mobile fleet and fixed facilities to maintain it.

1.15 Recovery Methods

The proposed processing strategy for the Carangas Project is based on a staged development of processing capacity over the LOM. During the initial operating period (Phase 1), Years 1 to 5, mineralised material will be treated through a single silver-lead-zinc process plant with a nominal design throughput of 22,000 tonnes per day (t/d) to produce a lead-silver concentrate and a zinc concentrate. Processing capacity is planned to increase during the expansion period (Phase 2), Years 6 to 8, to a nominal 44,000 t/d through the installation of a twinned 22,000 t/d process plant, duplicating the principal unit operations and associated infrastructure required for silver-lead and zinc concentrate production. In Year 9, one of the two process plants is planned to be converted (comminution and rougher flotation to be used) for the treatment of gold-bearing material and will operate as a dedicated 22,000 t/d Au processing facility after adding dedicated gold recovery equipment including cyanide leaching, counter-current decantation thickeners, Merrill Crowe and refining (Phase 3). One of the silver-lead-zinc plants will continue to operate as a 22,000 t/d concentrator. This dual-processing configuration is expected to continue until Year 16, with the gold plant converted back to a concentrator circuit for the remainder of the LOM. The phased approach provides flexibility to align processing capacity and flowsheet configuration with the mine production schedule and evolving feed characteristics.

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The crushing plant is designed at 75% availability, the lead-silver-zinc comminution and flotation circuit availability is designed at 91.3% and availability of the filter plants at 89%. The gold plant mill, flotation and leaching circuit availability is designed at 91.3%, with the gold room designed for availability of 52 weeks per year.

The process flowsheet for the project is based on preliminary metallurgical laboratory testing, as discussed in Section 13, and preliminary economic modelling. The selected unit operations are conventional technologies commonly applied in silver, lead, zinc, and gold processing plants of similar throughput, operating 24 hours per day and 365 days per year.

The proposed flowsheet includes the following process areas and unit operations:

· Crushing – primary crushing followed by stockpiling.
· Grinding – semi-autogenous grinding (SAG) mill, followed by ball milling with cyclone classification.
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· Silver-lead flotation – rougher, interstage thickening, concentrate regrind and three-stage cleaning.
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· Silver-lead concentrate handling – concentrate thickening and filtration.
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· Zinc flotation – rougher, concentrate regrind and three-stage cleaning.
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· Zinc concentrate handling – concentrate thickening and filtration.
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· Gold flotation – rougher flotation and thickening of rougher tailings.
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· Gold leaching – pre-leach thickening and cyanide leaching followed by counter-current decantation<br>(CCD).
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· Gold recovery –zinc precipitation of the clarified pregnant solution and smelting to produce doré.
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· Cyanide destruction and tailings management – tailings cyanide destruction, followed by tailings<br>thickening.
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Figure 1-2:      ConcentrateProcess Plant Flowsheet

Source: Ausenco, 2026

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Figure 1-3:      GoldProcess Plant Flowsheet

Source: Ausenco, 2026

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1.16 Project Infrastructure
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1.16.1 On-Site and Off-Site Infrastructure
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The Carangas Project will make use of current existing off-site infrastructure such as roads, transmission lines, and port facilities (at Iquique). Additional on-site infrastructure will be required to be built during the construction phase of the Project.

The on-site infrastructure for the Project includes site access roads, site roads and laydowns, WRSF, oxide stockpile, low-grade stockpile, gold zone stockpile, ROM stockpile, operations camp, auxiliary buildings, mine office and mine dry, truck shop, maintenance shop, and warehouse, mine rescue facilities, fuel storage and distribution facilities, power plant and electrical distribution, tailings storage facility (TSF) and water management structures, and an explosives magazine.

The off-site infrastructure required for the project includes National Highway 12 from the City of Oruro to the Town of Sabaya (existing), 36 km site access road from Sabaya to site (to be upgraded), and the 230 kV overhead transmission line from Pagador to Carangas (to be built).

1.16.2 Tailings Storage Facility (TSF)

A siting trade-off study was conducted among potential storage sites. The preferred deposition technology for the project is thickened slurry tailings. The preferred site is in the northwest corner of the large valley, upstream of the proposed open pit.

The TSF design has a capacity of 251 Mt. However, an additional 40 Mt of capacity is possible if the LOM plan is extended through additional drilling, mineral resource and reserve identification, and positive technical and economic studies. Thickened tailings from the process plant will be pumped to the TSF via a slurry pipeline and discharged off the crest of the embankment, creating a beach and a decant pond at the northwest end of the facility.

The TSF includes planning and management of slurry tailings, a starter facility, surface water management, an underdrain system, instrumentation, water quality monitoring systems, and a cover system at closure. The cover system will consist of a mixture of waste rock, tailings, alluvium, and topsoil with natural vegetation, designed to reduce erosion and limit surface water infiltration.

The PEA-level design has been developed in accordance with Canadian Dam Association guidelines and applicable Bolivian standards to ensure a safe and environmentally acceptable facility.

1.17 Market Studies and Contracts

No market studies or product valuations were completed as part of this study. Market price assumptions were based on a review of public information, industry consensus, standard practice, and specific information from comparable operations.

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Silver-lead and zinc concentrates are widely traded and can be marketed directly from producer to smelter or via third-party concentrate trading entities. It is assumed that the concentrate contains negligible deleterious elements that would impact marketability.

The market for gold doré is widely traded and can be marketed domestically or internationally with significant optionality regarding the final customer. It is assumed that the doré contains negligible deleterious elements that would impact marketability.

A marketing study was not conducted to determine indicative treatment and refining terms. Marketing, refining, and transportation costs, along with payability terms, were informed by preliminary discussions with a major commodity trading company and information from comparable recent studies.

1.18 Environmental, permitting and social considerations
1.18.1 Environmental Considerations
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1.18.1.1 Baseline and Supporting Studies
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Select baseline and supporting studies have been completed by Tierralta S.R.L (Tierralta) on behalf of NPM. These studies were completed in 2024 and include:

· Watershed Characterization
· Water Quality and Hydrology
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· Biodiversity Assessment
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· Ecosystem Functions
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· Ambient Air Quality.
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1.18.1.2 Environmental Monitoring
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Environmental monitoring for mining projects in Bolivia is regulated by the Ministerio de Medio Ambiente y Agua (Ministry of Environment and Water, MMAyA), through national and departmental environmental authorities, under the General Environmental Law.

Environmental monitoring will occur throughout the mine life, as indicated by national, departmental, and local regulations. Further post-closure environmental monitoring will align with the requirements for mine closure in Bolivia.

The proposed conceptual methods of waste rock and tailings management are currently considered by regulators and standards associations as generally acceptable practices. Additional testwork will be required to determine the degree of potentially acid-generating (PAG) and non-acid generation (NAG).

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1.18.2 Closure and Reclamation Considerations
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Closure of the Project will be primarily regulated by the Viceministerio de Medio Ambiente (Deputy Minister of the Environment), part of MMAyA, the Autoridad Jurisdiccional AdministrativaMinera (Mining Administrative Jurisdictional Authority, AJAM), and local governments. The legal framework for mine closure in Bolivia is governed by three key instruments: Environmental Law 1333, the Reglamento Ambiental para Actividades Mineras (Environmental Regulations for Mining Activities, RAAM), and the environmental licensing system (RPCA/EIA). As part of the permitting stage, the Project must submit a Closure Plan along with the Environmental Impact Assessment (EEIA) application. The Plan must include objectives, closure program details, rehabilitation measures, contaminant control measures, and post-closure monitoring. The Closure Plan will be prepared in parallel with mine facility designs, incorporate baseline information studies, and other operational and long-term planning efforts.

The Project will be closed in two phases: the active closure phase and the passive post-closure phase. During the active closure phase, closure reclamation activities for the mining, process plant, infrastructure, and TSF will take place. Environmental monitoring is assumed to be conducted during the reclamation and post-closure phases. The requirement for water treatment will be further assessed based on water balance considerations and geochemistry source term studies to be completed as the Project progresses through advanced studies. Treatment of water utilizing passive treatment systems will be a consideration during the passive post-closure phase.

1.18.3 Permitting Considerations

The Property consists of two mining areas, Granville and Colapso, granted by AJAM, the Bolivian mining authority. The two mining areas are currently advancing through the transition process from PEL to Administrative Mining Contracts (AMC) in accordance with Bolivian mining regulations.

During 2025, New Pacific Metals Corp. (NPM) initiated and advanced this regulatory transition process. Work and Investment Plans for the Granville and Colapso mining areas were prepared and submitted to the relevant authorities and subsequently reviewed with SERGEOMIN and AJAM as part of the AMC evaluation and approval process.

Upon approval of the AMCs by AJAM, the Company intends to apply for environmental categorization and initiate preparation of the EEIA required by future project environmental permitting for mine development.

New Pacific does not have direct title in the Carangas Project, as the Carangas Project is located within 50 km from international border where foreign companies or foreigners are not permitted to have ownership of land and right of mineral. Rather, Granville remains the holder of all licenses, permits and rights granted to it by Bolivian authorities and all operations on the Carangas Project are undertaken by Granville. New Pacific’s interest in the Carangas Project is as a result of the Mining Association Contract (MAC) entered into by wholly-owned subsidiaries of the Company and Granville. New Pacific is permitted to enter into the MAC with Granville under Bolivian Law No. 535 on Mining and Metallurgy. In April 2023, Granville initiated the corresponding process to request the authorization and registration of the MAC before AJAM. The Company’s indirect ownership in the Carangas Project will be certified as legally complaint with Bolivian laws related to the Frontier Area once the MAC is approved by the AJAM, registered in the Mining Registry, and published in the Mining Gazette (Bolivia).

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1.18.4 Social Considerations
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The Carangas Project is located within the Carangas TIOC (Indigenous, Native, and Peasant Territory). NPM has adopted a proactive and progressive engagement strategy focused on long-term collaboration, transparency, and responsible development in accordance with Bolivian regulations and international mining practices.

Throughout 2024, 2025, and early 2026, NPM implemented an extensive community engagement program designed to strengthen relationships with local rightsholders and facilitate the advancement of the Project. Engagement activities included formal and informal meetings with community members and authorities, technical workshops, educational outreach programs, stakeholder mapping, sociocultural baseline studies, and communication campaigns tailored to local customs and traditions. Additional support initiatives included veterinary assistance and technical support programs for local camelid livestock producers, as well as participation in civic, social, and cultural activities within the community.

On January 26, 2026, NPM and the Carangas Community executed a formal Framework Agreement regulating the relationship between the Project and the Community. This agreement represents the most significant advancement achieved to date in obtaining the social permits required for project development and establishes the principal governance instrument currently in force between both parties.

On July 6, 2026, NPM successfully completed a prior consultation to obtain community and other regional stakeholders’ consent as part of the requirements for the conversion of the Carangas ELs to AMCs. The application process for the ELs to AMCs conversion started in 2025 and all required documents have been submitted to the Bolivia’s Ministry of Mining and Metallurgy since then. With the recently completed prior consultation, the Company expects that the remaining administrative work and legislative approval of the conversion of the ELs to AMCs could be completed within six months or less.

The project as it is envisioned will require the movement of the town of Carangas and prior to the commencement of open pit mining activities. This has been discussed with the community and a preferred location selected. An allowance has been made in the capital estimate to move the town.

Once the AMCs are obtained, the Company will start the application to obtain its environmental categorization as a proposed open pit operation from Bolivia’s Ministry of Environment and Water, formally commencing the EEIA process. Work will also commence on gathering baseline environmental and social data as well as other associated technical work to fulfill the requirements of the EEIA. It is expected that this work will be completed by the end of 2027.

1.19 Capital and Operating Cost

The capital and operating costs described in this PEA are based on open pit mining operations for the Carangas project. The process plant is designed to treat 22,000 t/d (8.0 Mt/a) of mineralized material during the initial phase, and 44,000 t/d (16.0 Mt/a) during the expansion phase, over a mine life of 19 years.

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1.19.1 Capital Cost Estimate
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The capital cost estimate was developed in Q2 2026 to target a level of accuracy of -30% to +50%, which aligns with an Association for the Advancement of Cost Engineering (AACE) International Class 5 level estimate. The estimate includes mining, processing, on-site infrastructure, off-site infrastructure, project indirects, project delivery, owners’ costs, and provisions. The total initial capital costs for the Project are estimated at US$644.5 million, including US$55.3 million of capitalized operating costs, and US$110.3 million of contingency. The total expansion capital costs are US$422.7 million and the total LOM sustaining costs are estimated at US$166.5 million, while the closure costs are estimated at US$149.8 million. The capital cost summary is presented in Table 1-3.

Table 1-3:      CapitalCost Summary

WBS Description Initial Capital Cost (US$M) Concentrate Plant Expansion Cost (US$M) Gold Plant Expansion Cost (US$M) Sustaining Cost (US$M) Total Cost (US$M)
1000 Mining $3.4 - - $14.6 $18.1
2000 Concentrate Plant $202.1 $202.1 - - $404.3
3000 Gold plant - - $42.9 - $42.9
4000 On-site Infrastructure $65.8 $20.3 $0.8 $151.9 $238.8
5000 Off-site Infrastructure $114.5 - - - $114.5
Total Direct Costs $385.9 $222.5 $43.8 $166.5 $818.6
6000 Project Preliminaries $31.7 $16.7 $3.3 - $51.7
7000 Project Delivery $45.6 $38.9 $7.9 - $92.4
8000 Owner’s Costs $15.7 - - - $15.7
Total Indirect Costs $93.0 $55.6 $11.2 - $159.8
Total Direct + Indirect Costs $478.9 $278.0 $54.9 $166.5 $978.3
Contingency $110.3 $74.6 $15.1 - $200.0
Subtotal Capital Cost $589.2 $352.6 $70.0 $166.5 $1,178.3
Capitalized Opex $55.3 - - - $55.3
Closure Costs - - - $149.8 $149.8
Total Capital Cost $644.5 $352.6 $70.0 $316.3 $1,383.4

Note: Totals may not match due to rounding.

The capital cost estimate was developed in Q2 2026 United States dollars (USD). The estimate is based on budgetary quotations for equipment from recent advanced studies and execution projects, supplemented with Ausenco’s in-house database, and informed by Ausenco’s experience from similar operations in South America.

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The following data were used as the basis of estimate:

· Mining schedules (mining is assumed to be done by a contractor who will provide equipment and maintenance<br>facilities)
· Engineering design by Ausenco, including but not limited to design criteria, equipment lists, and material<br>take-offs (MTOs)
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· Budgetary equipment quotes from similar recently completed advanced studies and execution projects, escalated<br>to Q2 2026
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· Additional data such as lang factors and indirect costs from similar recently completed studies and projects.
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The estimate also adhered to these parameters:

· No allowance was made for exchange rate fluctuations
· No additional escalation was added to the final estimate
--- ---
· No price growth or design allowance was included.
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1.19.2 Operating Cost Estimate
--- ---

The total operating costs for the Project are estimated at US$16.93/t or US$4,258.4 million over the 19-year mine life. These operating costs do not include pre-production operating costs. A summary of operating costs is presented in Table 1-4.

Table 1-4:      OperatingCost Summary

Cost Area Total (US$M) US$/t % of Total
Mining 2,120.2 8.43 50
Process 1,791.2 7.12 42
G&A 347.0 1.38 8
Total 4,258.4 16.93 100

Mine operating costs were built up from first principles assuming a contracted mining operator and an owner managed and operated mine site.

The following was used to determine the project’s LOM process operating costs in agreement with the cost definition and estimate methodologies outlined below. This basis considers the development of a process plant designed to treat 22,000 t/d of mineralized material in the initial phase, and 44,000 t/d in the expansion phases. Process unit operations were benchmarked against similar or comparable processing plants to ensure accuracy of cost estimates.

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Assumptions made in developing the process operating cost estimate are listed below:

· Mill production is designed to treat 22,000 t/d of mineralized material in the initial phase, and 44,000<br>t/d in the expansion phases.
· Process plant operating costs are calculated based on labor, power consumption, and process and maintenance<br>consumables.
--- ---
· Off-site gold refining, insurance, and transportation costs are excluded, as they are included elsewhere<br>in the financial model.
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· Labor rates were sourced from recent execution projects in the region.
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· Workforce will be comprised of local and regional workers.
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· Management and administrative staff will be on a 5/2 rotation (5 days in, 2 days out), whereas process<br>and maintenance staff will be on a 14/7 rotation (14 days in, 7 days out).
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· Management and administrative staff who are not required to be on site will be based out of La Paz or<br>will work remotely.
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· General and administration (G&A) costs were baselined against previous regional project experience.
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· No factor for spare parts has been applied to adjust for consumption of fewer spare parts in early years<br>of operation.
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· Grinding media consumption rates have been estimated based on the mill feed characteristics.
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· Reagent consumption rates have been estimated based on the metallurgical testwork results.
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· Reagents and consumable prices were obtained via recent regional quotations with location-specific freight<br>considerations.
--- ---
· Mobile equipment cost includes for fuel, maintenance, and lease price for the equipment.
--- ---
· A unit power cost of US$0.06/kWh supplied to NPM by Empresa Nacional de Electricidad (ENDE) Transmisión<br>S.A.
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· A unit fuel cost of US$5.34/gal (US$1.41/L), based on current regional fuel prices at the Project site<br>by NPM.
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1.20 Economic Analysis
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1.20.1 Economic Summary
--- ---

Readers are cautioned that the PEA is preliminary in nature, that it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the preliminary economic assessment will be realized. The capital and operating cost estimates were developed in Q2 2026 to target a level of accuracy of -30% to +50%, which aligns with an AACE International Class 5 level estimate. The capital cost estimate includes a 23% contingency on the initial capital costs.

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The PEA has been evaluated using a discounted cashflow (DCF) analysis. Cash inflow consists of annual revenue projections for the Project. Cash outflows such as capital costs, operating costs, taxes, and royalties are subtracted from the inflows to arrive at the annual cashflow projections.

The post-tax NPV at a 5% discount rate (NPV5%), is US$2.65 billion with a post-tax internal rate of return (IRR) of 35.9%, and an initial payback of 2.4 years. These economic results utilize base-case prices of US$3,400/oz Au, US$45.00/oz Ag, US$0.90/lb Pb, and US$1.20/lb Zn.

1.20.2 Sensitivity Analysis

A sensitivity analysis was conducted on the base-case pre-tax and post-tax NPV5% and IRR of the project using the following variables: metal prices, discount rate, total operating costs, initial capital costs, recovery, and head grade. The sensitivity analysis reveals that the project is most sensitive to changes in recovery, head grade and commodity prices, and less sensitive to operating costs, and initial capital costs.

1.21 Conclusions & Recommendations

The study highlights positive economic results based on a conceptual design. Further exploration work is recommended to evaluate the Project’s mineral potential. Additional field work, laboratory testwork, and analysis are required prior to advancing to a PFS. It is also recommended that the Company initiates environmental baseline studies and engages a permitting consultant along with ramping up engagement with government, local groups, local communities, and regional infrastructure owners. The recommended work is estimated to cost US$37.5 million and is summarized in Table 1-5.

Table 1-5:      RecommendedWork Programs

Program Component Estimated Total Cost (US$M)
Drilling 12.0
Geology and Exploration 0.7
Mining and Mine Open Pit Geotechnical Program 3.8
Metallurgical Testing 2.5
Recovery Methods 3.5
Infrastructure Geotechnical 1.4
TSF and WRSF Analyses and Studies 0.6
Hydrogeological - Stage 1 1.5
Hydrogeological - Stage 2 2.0
Hydrological 1.0
Geochemical 2.5
Environmental 6.0
Total 37.5
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2 Introduction
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2.1 Introduction
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Ausenco Engineering Canada ULC, and Ausenco sustainability (collectively, Ausenco) has prepared a preliminary economic assessment (PEA) and associated technical report for New Pacific Metals Corp. (NPM) on the Carangas Project, located in the Carangas District, situated on the western part of the Oruro Department, in Bolivia. The technical report was prepared in accordance with the Canadian disclosure requirements of National Instrument 43-101 – Standards of Disclosure for Mineral Projects (NI 43-101) and Form 43-101 Form 1 – Technical Report.

The responsibilities of the engineering consultants are as follows:

· Ausenco was responsible for the process plant, infrastructure, tailings, waste rock, water management,<br>environmental, permitting, and community, cost estimation, financial modelling, and overall report compilation.
· Moose Mountain Technical Services (MMTS) was responsible for mining, stockpiles, haul roads, and mining-related<br>cost estimation.
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· SLR Advisory Brasil Ltda (SLR) was responsible for the property, geology, exploration, drilling, and sampling<br>in addition to the mineral resource estimation.
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· JJ Metallurgical Services Inc. was responsible for the metallurgical testwork and recoveries.
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2.2 Qualified Persons
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The Qualified Persons for the report are listed in Table 2-1. By virtue of their education, experience and professional association membership, they are considered Qualified Persons as defined by NI 43-101.

Table 2-1:      ReportContributors

Qualified Person Professional Designation Position Employer Independent of NPM
Kevin Murray P.Eng. Principle Process Engineer Ausenco Yes
Scott C. Elfen PE, P.Eng. SME and Global Lead (Geotechnical) Ausenco Yes
James Millard P.Geo. Director, Strategic Projects Ausenco Yes
Jinxing Ji P.Eng. Principle Consultant JJ Metallurgical Services Yes
Anderson G. Candido FAusIMM Principle Resource Geologist SLR Yes
Marc Schulte P.Eng. Mining Engineer MMTS Yes
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2.3 Terms of Reference
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The report supports disclosures by NPM in a press release dated July 16, 2026, and entitled "New Pacific Metals Reports Results of Updated Carangas Preliminary Economic Assessment:Post-tax $2.65 Billion NPV (5%) and 35.9% IRR; 339.0 Million Oz of Silver Equivalent Produced.”

All measurement units used in this report are SI/metric units unless otherwise noted. Currency is expressed in United States dollars (US$ or USD) unless otherwise noted.

Mineral resources and mineral reserves are reported in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and Mineral Reserves (CIM, 2014) and the CIM Estimation of Mineral Resources and Mineral Reserves Best Practice Guidelines (CIM, 2019).

2.4 Site Visits and Scope of Personal Inspection
2.4.1 Personal Inspection Details by Anderson G. Candido, FAusIMM
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Mr. Anderson Candido, SLR’s Geological QP, visited the site March 27–30, 2023. The purpose of the visit was to inspect key aspects of the geology, including drill collar locations, mineralized outcrops, historical artisanal workings, drill pad area, drill core intervals, QA/QC protocols, and core storage facilities. Additionally, the QP has had several technical discussions with the Carangas team to verify the overall geological procedures and the adopted strategy for the exploration program.

2.4.2 Personal Inspection Details by Scott Elfen, PE, P.Eng.

Mr. Scott C. Elfen, Ausenco’s Geotechnical QP, visited the site March 16–17, 2026. The purpose of the visit was to assess the site’s topography, surface geotechnical conditions, and water features, and their amenability to support tailings and waste rock structures, and to review the open pit geometry and drill core to develop the future geomechanical program. Additionally, the QP assessed site access, the local community, and the overall layout of mining related infrastructure.

2.4.3 Personal Inspection Details by Marc Schulte, P.Eng.

Mr. Marc Schulte, MMTS’s Mining QP, visited the Carangas site March 16–17, 2026. On the visit he assessed the topography and physiography covering planned open pit, mill feed stockpiles, waste storage areas, and plant and office areas. The QP also traversed the site access from local communities as well as larger population centres in western Bolivia, and assessed any potential restrictions (narrow roads, bridges, etc.). The QP viewed drill core from both the upper silver zone and the lower gold zone, as well as barren areas

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2.4.4 Personal Inspection Details by Jinxing Ji, P.Eng.
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Mr. Jinxing Ji, JJ Metallurgical Services’ Metallurgical QP visited the Carangas Site on May 21–23, 2022. During the site visit, he reviewed the deposit area, discussed aspects of the deposit with the geologists, and compiled samples for metallurgical testing.

2.4.5 Personal Inspection Details by Kevin Murray, P.Eng.

Mr. Kevin Murray, Ausenco’s Process, Infrastructure, and Economics QP, has not visited the Carangas site. Due to the early stage of the Project development, and conceptual process and infrastructure design, the QP determined that a site inspection was not required.

2.4.6 Personal Inspection Details by James Millard, P.Geo.

Mr. James Millard, Ausenco’s Environmental, Permitting, Social or Community Impact QP has not visited the Carangas site. Due to the early stage of the Project development and environmental work, as well as work completed by Tierralta S.R.L. (an environmental consultancy based in La Paz, Bolivia) the QP deemed a site inspection was not necessary.

2.5 Sources of Information
2.5.1 Information and Data
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Sources of information are listed in Section 27, with the sources provided by NPM regarding environmental studies listed in Section 3.

The QPs have not independently reviewed the current status or future prospects of mineral exploration and/or mining permits held or required by NPM for work on the Project. The QPs have fully relied on information supplied by NPM.

2.5.2 Previous Technical Reports

The previous technical report was published on November 15, 2024, and had an effective date of September 5, 2024. It disclosed an MRE and the results of a PEA and was prepared by RPM Global for NPM.

2.6 Effective Dates

The report contains the following effective dates:

· Effective date of the Mineral Resource Estimate: March 31, 2026.
· Effective date of the Financial Model: July 16, 2026.
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· Effective date of the Report is: July 16, 2026.
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2.7 Currency, Units, Abbreviations and Definitions
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All units of measurement in this report are metric and all currencies are expressed in United States dollars (symbol: US$ or currency: USD) unless otherwise stated. Contained gold metal is expressed as troy ounces (oz), where 1 oz = 31.1035 g. All material tonnes are expressed as dry tonnes (t) unless stated otherwise. A list of abbreviations and acronyms is provided in Table 2-2 and units of measurement are listed in Table 2-3.

Table 2-2:      Abbreviationsand Acronyms

Abbreviation Description
AAC Autoridad Ambiental Competente (Competent Environmental Authority)
AACE Association for the Advancement of Cost Engineering International
AAS atomic absorption spectroscopy
ABC andesitic basalt
AET actual evapotranspiration
AJAM Autoridad Jurisdiccional Administrativa Minera (Mining Administrative Jurisdictional Authority)
ALS ALS Laboratories
AMC Administrative Mining Contract
Au gold
Az azimuth
BIF banded iron formation
BBWi bond ball mill work index
CAD:USD Canadian-American exchange rate
CCD counter-current decantation
CDA Canadian Dam Association
CDN CDN Resource Laboratories of Langley, British Columbia, Canada
CDSF Co-disposal storage facility
CIM Canadian Institute of Mining, Metallurgy and Petroleum
CIM Definition Standards CIM Definition Standards for Mineral Resources and Mineral Reserves 2014
CIP carbon-in-pulp
CNPV Census Nacional de Población y Vivienda (National Population and Housing Census)
CNWAD Weak Acid Disociable Cyanide
COD chemical oxygen demand
CoG cut-off grade
COMSUR Compañía Minera del Sur S.A.
CRM certified reference material
CV coefficient of variation
CWi Bond Crusher Work Index
DCF discounted cashflow
DCIP direct current resistivity and induced polarization
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Abbreviation Description
--- ---
DDH diamond drill hole
DTH down-the-hole
EDGM Earthquake Design Ground Motion
EEIA Estudio de Evaluación de Impacto Ambiental (Environmental Impact Assessment Study)
E-GRG extended gravity recoverable gold
EIA Environmental Impact Assessment
EM electromagnetic
EMP Environmental Management Plan
ENDE Empresa Nacional de Electricidad (National Electricity Company)
EPCM Engineering, Procurement, and Construction Management
FA fire assay
FET federal excise tax
FoS Factor of safety
FS feasibility study
G&A general and administration
GPR gross production royalty
GPS global positioning system
GQCV greenstone-hosted quartz-carbonate vein deposits
Granville Minera Granville S.R.L.
GRAV gravimetric finish method
HDPE High-Density Polyethylene
HU hydrographic units
ICP inductively coupled plasma
ICP-OES inductively coupled plasma - optical emission spectrometry
ID2 inverse distance squared
ID3 inverse distance cubed
IDF Inflow design flood
IEC International Electrotechnical Commission
IHF Indice de Hábitat Fluvial (Fluvial habitat index)
IOCG iron oxide copper gold
IP induced polarization
IPA Indice de Presión Antropogénica (Anthropogenic Pressure Index)
IRGS intrusion-related gold system
IRR Internal rate of return
ISO International Organization for Standardization
KCB Klohn Crippen Berger
LGSP Low-Grade Stockpile
LiDAR light detection and ranging
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Abbreviation Description
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Llicancabur Llicancabur Mining Ltda.
LOM Life of mine
LUP land use permit
MAC Mining Association Contract
MCE maximum credible earthquake
MCF mechanized cut and fill
MEL mechanical equipment list
MMAyA Ministerio de Medio Ambiente y Agua (Ministry of Environment and Water)
MMM Ministerio de Minería y Metalurgia (Ministry of Mining and Metallurgy)
MMTS Moose Mountain Technical Services
MRE mineral resource estimate
NaCN Sodium cyanide
NAD 83 North American Datum of 1983
NAG Non-acid generating
NI 43-101 National Instrument 43-101; Standards of Disclosure for Mineral Projects
NN nearest neighbour
NPM New Pacific Metals Corp.
NSP net smelter price
NSR net smelter return
NTS national topographic system
OK ordinary kriging
OXSP Oxide Stockpile
PAG Potential acid generation
Pb lead
PEA preliminary economic assessment
PEL Prospecting and Exploration Licenses
PFS prefeasibility study
PGA Peak ground acceleration
PGE platinum group elements
PMF Probable maximum flood
QA/QC quality assurance/quality control
QBR Qualitat del Bosc de Ribera (Riparian Forest Quality Index)
QP qualified person (as defined in National Instrument 43-101)
RAAM Reglamento Ambiental para Actividades Mineras<br><br> (Environmental Regulations for Mining Activities)
ROM run of mine
RPCA Reglamento para la Prevención y Control Ambiental<br><br> (Regulations for Environmental Prevention and Control)
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Abbreviation Description
--- ---
RPEEE Reasonable prospects for eventual economic extraction
RQD rock quality designation
RTK Real time kinematic
SAG semi-autogenous grinding
SCC Standards Council of Canada
SD standard deviation
Sd-BWI micro hardness or bond ball mill work index on SAG ground material
SEDEX sedimentary exhalative deposits
SERGEOMIN Servicio Geológico Minero (Geological Mining Service), Bolivia
SG specific gravity
SMBS sodium metabisulfite
SMU Selective mining unit
SPT Standard penetration test
TCC Total cash cost
Tierralta Tierralta S.R.L.
TIOC Territorio Indígena Originario Campesino (Native Indigenous Peasant Territory)
TMF tailings management facility
TSF tailings storage facility
UCF Undiscounted cash flow
UG underground
USZ Upper silver zone
UTM Universal Transverse Mercator coordinate system
UV ultraviolet
VLF-EM very low frequency electromagnetic
VMS volcanogenic massive sulfide
VSA Vacuum swing adsorption
WRSF Waste rock storage facility
XRF x-ray fluorescence
Zn zinc
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Table 2-3:      Unitsof Measurement

Abbreviation Description
% percent
% solids percent solids by weight
CAD Canadian dollar (currency)
C$ Canadian dollar (as symbol)
$/t dollars per metric ton
° angular degree
°C degree Celsius
μm micron (micrometer)
cm centimeter
cm^3^ cubic centimeter
ft foot (12 inches)
g gram
g/cm^3^ gram per cubic centimeter
g/L gram per liter
g/t gram per metric ton (tonne)
h hour (60 minutes)
ha hectare
kg kilogram
kg/t kilogram per tonne
km kilometer
km^2^ square kilometer
kW kilowatt
kWh/t kilowatt-hour per tonne
L liter
lb pound
m, m^2^, m^3^ meter, square meter, cubic meter
M million
Ma million years (annum)
masl meters above sea level
mm millimeter
Moz million (troy) ounces
Mt million tonnes
MW megawatt
oz troy ounce
oz/t ounce (troy) per tonne
oz/ton ounce (troy) per short ton (2,000 lbs)
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Abbreviation Description
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ppb parts per billion
ppm parts per million
t metric tonne (1,000 kg)
ton short ton (2,000 lbs)
t/d tonnes per day
USD US dollars (currency)
US$ US dollar (as symbol)
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3 Reliance on Other Experts
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3.1 Introduction
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The QPs have relied upon the following other expert reports, which provided information regarding environmental, permitting, and social and community impacts for sections of this Report.

3.2 Environmental, Permitting, Closure, and Social and Community Impacts

The QPs have fully relied upon information supplied by NPM and experts retained by NPM for information related to environmental, permitting, and social and community impacts as follows:

· Tierralta<br> S.R.L, (2024): Environmental Study with a Multisectoral Approach - Carangas Mining Project.<br> Prepared for New Pacific Metals Corp., December 2024, 93 pp.

This information is used in Sections 1.18, 20, 25.13, and 26.8 of the Report.

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4 Property Description and Location
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4.1 Property Description
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The Carangas Property is located in the Carangas region in the western portion of the Department of Oruro, Bolivia, approximately 190 km from the city of Oruro. The coordinates of the center of the Property are 7,906,871 Northing and 541,116 Easting (WGS84, UTM Zone 19S), corresponding to latitude 18°55'48.05" S and longitude 68°36'34.25" W. The average altitude is approximately 3,950 meters above sea level (masl). The Property has a total area of 39 km² (3,900 ha). The location of the Property is shown in Figure 4-1.

Figure 4-1:          CarangasProject General Location Plan

Source: NPM, 2023

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The Carangas Project is located in the Oruro region of Bolivia and is covered by two Prospecting and Exploration Licenses (PELs) issued by the Mining Administrative Jurisdictional Authority (AJAM) under Bolivian mining law. These licenses were granted to Granville, and NPM has a Mining Association Contract (MAC) with Granville to conduct the geological and mining works related to these PELs. In Bolivia, Mineral Resources belong to the State, and companies obtain the right to explore or mine through licenses or administrative mining contracts. Exploration licenses are typically granted for a fixed term and must be kept in good standing through annual payments, filing of required reports, and compliance with environmental regulations. The Carangas licenses are current, and all legal and administrative obligations have been met.

Surface rights in Bolivia are separate from mineral rights. While a PEL grants the right to explore for minerals, it does not grant ownership or automatic access to the land surface. The ground within the Carangas area belongs to local communities, and access must be negotiated directly with them. NPM has secured the necessary agreements to conduct exploration activities, including the construction of roads, drill pads, and camps. Future mine development, including the construction of a plant, tailings storage facility (TSF), and other permanent structures, would require new land use agreements with these same community landowners.

Water rights and royalties are also governed by Bolivian law. Water use for exploration or mining requires separate authorization, which defines permitted volumes and sources. The project area includes perennial streams and access to the Todos Santos River, however, long-term water use for mining purposes would need formal approval through the national and regional water authorities. Bolivia also has a State royalty system that applies to mineral production. Royalty rates vary by commodity and are calculated based on the value of the minerals sold. While royalties are not triggered during exploration, they will apply if a project advances to mining. Based on current information, the mineral tenure for the Carangas Project is secure, and there are no known legal or jurisdictional issues that would prevent continued exploration or preparation of Mineral Resource estimates.

4.2 Project Ownership

The Carangas Project consists of two PELs granted to Granville by AJAM. In April 2021, NPM entered into a Mining Association Contract with Granville to acquire a 98% interest in the Carangas Project. Granville is wholly owned by Bolivian nationals and retains legal title to the exploration licences under Bolivian mining regulations.

Under the Mining Association Contract structure, Granville remains the legal holder of the PELs, while NPM participates in the Project through contractual rights that allow it to fund and carry out exploration, development, and related technical activities in accordance with the terms of the agreement. The Company is responsible for funding project expenditures, including exploration, development, and potential production activities.

The agreement has a term of 30 years and is renewable for an additional 15 years. It provides NPM with effective control over the advancement of the Project through its financing and operational decision-making rights, while maintaining compliance with the Bolivian legal framework governing mineral title ownership and exploration rights.

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The pit design for the deeper gold zone requires mining of waste (waste stripping) on mining concessions in the southern portion of the planned open pit that do not belong to the Company. These concessions include approximately 1.85% of the mineral resources that have been included in the economic analysis for this PEA. These concessions are held by the state of Bolivia and are not currently available for tenure. Although the Company is actively working with the Bolivian government to obtain them, failure to do so, or to enter into a mining agreement on them could cause the Company to reevaluate the pit design and the outcome of this PEA.

4.3 Mineral Tenure

The Property consists of two PELs. Details for the concessions are provided in Table 4-1 and Figure 4-2. The concessions are currently transitioning from PELs to an Administrative Mining Contract (AMC), which secures the rights to operate the areas during this process and to advance the Project. The concessions are current as of the effective date of this Report, and will remain valid until the AMC is granted.

According to Bolivian regulations, as the Property is located within 50 km of international borders, foreign companies or foreigners are not permitted to have ownership of the land and right of mineral, Granville remains the holder of all licenses, permits, and rights granted to it by Bolivian authorities. To the extent known, there are no other royalties, back-in rights, payments, or other agreements and encumbrances to which the Property is subject.

The Carangas Project comprises two areas: GRANVILLE—created by consolidating the GRANVILLE I and GRANVILLE II areas—and COLAPSO. Granville was notified of the PEL for GRANVILLE II on April 15, 2021, and of the PELs for GRANVILLE I and COLAPSO on November 21, 2022. Each PEL was granted for a five-year term from its notification date. In April 2023, Granville exercised its preferential right and submitted the corresponding Administrative Mining Contract applications over these areas. The applications remain in process with AJAM.

Table 4-1:          MiningRights of the Carangas Property

Area Code Name Concession Type Size in km^2^ (ha) Title holder
1800001 COLAPSO PEL-AMC* 32.75<br> (3257) Minera<br> Granville S.R.L.
1800000 GRANVILLE PEL-AMC* 6.25<br> (625) Minera<br> Granville S.R.L.

*Transitioning from PEL to AMC. Source: NPM, 2026

The QP is not aware of any environmental liabilities on the Property. NPM has all the required permits to conduct the proposed work on the Property. The QP is not aware of any other significant factors and risks that may affect access, title, or the right or ability to perform the proposed work program on the Property.

NPM and Granville has the right to:

· Carry<br> out mining exploration and prospecting activities in the mining area indicated in the licenses<br> for a specific term.
· Commercialize<br> the eventual mineral production from the exploration activities.
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· Exercise<br> the "Preferential Right", which allows the holder to request the signing of an<br> AMC on the mining area preferentially over any other interested parties.
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· Exercise<br> the rights of passage, which allow transit through the Bolivian state land and/or neighbouring<br> properties to access the holder’s license area under prior agreement with the landowner<br> and to build paths, roads, bridges, pipelines, aqueducts, power lines, railway lines, and<br> install the necessary basic services, at its own expense and cost.
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To maintain the PELs in good standing, NPM and Granville shall:

· Commence<br> exploration and prospecting activities within one year from the date of the license grant.
· Not<br> suspend activities for more than one year without a justifiable reason.
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· Deliver<br> reports each semester on the progress of activities to AJAM.
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· Pay<br> the corresponding mining patent fees yearly in advance according to applicable Bolivian laws.
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· Obtain<br> the required environmental license before conducting prospecting and exploration activities<br> in the area.
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· Follow<br> the current exploration/mining Bolivian legislation and regulations.
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Figure 4-2:          Locationof Carangas Mineral Rights

Source: NPM, 2026

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4.4 Property Agreements
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NPM entered into the Mining Association Contract (MAC) with Granville, establishing a joint venture arrangement for the development of permitted mining activities in accordance with applicable Bolivian laws. The joint venture grants NPM and Granville the opportunity to conduct mining activities within the mining area pursuant to the terms and conditions of the MAC.

Terms of the joint venture were disclosed by NPM in its management discussion and analysis for the three and nine months ended March 31, 2022, as follows (New Pacific SEDAR issuer profile – MD&A, May 11, 2022):

“InApril 2021, the Company signed an agreement with a private Bolivian company to acquire a 98% interest in the Carangas Project. Theproject is located approximately 190 km southwest of the city of Oruro and within 50 km from Bolivia’s border with Chile. The privateBolivian company is 100% owned by Bolivian nationals and holds title to two exploration licenses covering an area of 6.25 km^2^.Under the agreement, the Company is required to cover 100% of the future expenditures on exploration, mining, development, and productionactivities for the project. The agreement has a term of 30 years and is renewable for another 15 years.”

To maintain the Property in good standing, NPM and Granville must comply with the agreement in relation to the development of the authorized mining activities.

4.5 Project Risks and Uncertainties

Although the Carangas Project holds valid PELs issued by AJAM, there are several risks and uncertainties that relate to property ownership, surface access, and permitting. The mineral rights are held by Granville, a Bolivian company, because foreign companies cannot hold mineral titles within 50 km of an international border. NPM participates in the Project through a MAC, which grants it the right to explore and develop the Property but does not convey direct title ownership. This structure places responsibility for maintaining the licenses and meeting legal requirements on Granville, creating a dependency on a third-party title holder.

Surface rights at Carangas are held by the local communities, not by the license holder. While permissions have been secured for exploration activities at the Project (e.g., drilling, road building, and camp facilities), future project stages, including construction of a mine, processing plant, TSF, waste dumps, and water and power infrastructure, will require new agreements. On January 26, 2026, NPM and the Carangas community executed a formal Framework Agreement regulating the relationship between the Project and the community. This agreement represents the most significant advancement achieved to date in obtaining the social permits required for project development and establishes the principal governance instrument currently in force between both parties. Any changes in community relationships, leadership, or expectations could affect NPM’s ability to expand exploration or advance toward development. Community concerns may also be heightened by the Project’s proximity to the Carangas town and culturally sensitive areas, such as the colonial era church located near one of the potential TSF sites, which may influence approval timelines or project design requirements.

There are also uncertainties related to environmental and water-use permitting. To develop the Project beyond exploration, NPM must secure Environmental Impact Assessment (EEIA) approval and undergo public consultation. Given how close the town is to the proposed pit and infrastructure footprint, environmental and social concerns could delay or complicate the permitting process. Water rights present another risk, as long-term operational water supply from the Todos Santos River will require formal authorization from the relevant authorities. Although there are no known environmental liabilities on the Property, changes in regulation or new findings during baseline studies could impose additional requirements or constraints.

Overall, the Carangas Project currently holds the rights required for exploration. Still, property ownership structure, reliance on community surface access, future environmental approvals, water-use authorizations, and the need to maintain social license all represent foreseeable uncertainties that could influence the Project’s ability to progress beyond the current stage.

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5 Accessibility, Climate, Local Resources, Infrastructure and Physiography
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5.1 Physiography and Climate
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The Project area is in Bolivia's Western Andes Cordillera, a rugged mountain region with elevations ranging from 4,000 to 5,000 masl. Within the Project area, elevations range from 4,074 masl at the top of the West Dome to 3,904 masl at the Carangas Stream between the West Dome and the East Dome.

Vegetation on the Property consists of low grasses and shrubs. The climate of the Western Cordillera is cool and dry, especially in the winter months. In the Carangas area, high temperatures range from 12.4°C in July to 19.2°C in October, and the low temperatures range from -3.5°C in July to 3.8°C in January (Table 5-1). Rainfall in the area is sparse and ranges from 2 mm in June to 162 mm in January. The local climate does not limit the length of the operating season.

Table 5-1:          Weatherof Carangas Region

Jan. Feb. Mar. Apr. May Jun. Jul. Aug. Sep. Oct. Nov. Dec.
Avg.<br> Temperature °C 9.2°C 9°C 8.7°C 7.8°C 5.5°C 4.5°C 3.8°C 5.3°C 7.2°C 9°C 10.1°C 10.4°C
Min.<br> Temperature °C 3.8°C 3.8°C 2.4°C 0.3°C -2.4°C -3°C -3.5°C -2.9°C -1.5°C 0.2°C 1.2°C 3.3°C
Max.<br> Temperature °C 15.4°C 15.1°C 15.4°C 15.7°C 14.3°C 13.1°C 12.4°C 14.2°C 16.3°C 18.1°C 19.2°C 18.3°C
Precipitation<br> (mm) 162 138 81 21 3 2 4 6 6 12 22 82
Humidity<br> (%) 59% 65% 59% 40% 22% 16% 17% 17% 19% 21% 22% 37%
Rainy<br> days (d) 15 13 11 4 1 0 1 1 1 2 3 9
Avg.<br> Sun hours (hours) 8.6 7.9 8.7 9.8 10 9.8 9.8 10.2 10.7 11.1 11.5 10.6

Source: climate-data.org, 2023

5.2 Accessibility

The Project area is accessible by vehicle from Oruro, with approximately 190 km of paved road on National Highway 12 leading to the town of Sabaya, followed by 35 km on a flat gravel road from Sabaya to Carangas (Figure 5-1).

The closest major population centre is Oruro, with a population of approximately 260,000 people. Some small farming and grazing communities (pueblos) are scattered throughout the region, the closest being Carangas, situated between the two prominent hills (the West Dome and the East Dome) on the Property. The official population of the municipality of Carangas is 1,130, according to the 2024 Census; however, only a small number of people reside in the community on a regular basis, with the majority living in Oruro. Oruro has a long history and culture of mining dating back to the colonial era, and a sufficient supply of skilled mining talent will be available once the Project evolves into mining production.

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Figure 5-1:          CarangasProject General Location Plan

Source: NPM, 2023

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5.3 Local Resources and Infrastructure
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Two small local streams run through the Property, both with a flow rate of approximately 20 L/s each in the dry season. This water supply was more than sufficient for the Project's water consumption during the exploration and drilling stage. Approximately 5 km south of the Property, the streams join the larger Todos Santos River near the community of La Rivera, which flows at more than 1,000 L/s in the dry season and could provide an adequate alternative water supply for future mining stages if required. A 220 kV single-phase electric power line is available from Oruro via a 250 km-long transmission line that runs along National Highway 12. A three-phase industrial power line is available approximately 5 km to the south of the Property near the community of La Rivera.

The Carangas pueblo has essential facilities, including a medical clinic, school, church, soccer field, municipal offices, and mobile telecommunications towers.

The NPM has set up its exploration camp facilities at Carangas, including exploration offices, accommodation and dining, a core shack for logging and sampling, and storage. Most supplies for the Project are transported by truck from Oruro and La Paz.

The PELs grant the NPM and Granville only the right of exploration and small-scale mining. Surface rights belong to the local communities. The NPM has obtained permission from local communities to build drill roads, pads, and other exploration infrastructures during the exploration stage. Agreements or permissions need to be secured from the local communities to build the mine and other facilities, such as a processing plant, tailing and waste storage, as well as offices and accommodations in the vast open area surrounding the Property when it moves to the mining production stage.

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6 History
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6.1 Regional History
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The Property has a long mining history, dating back to the Spanish colonial period in the mid-1500s and continuing intermittently until the 20th century. In the 1980s, COMSUR was active in the area and collected 350 samples from surface dumps and underground channels.

From 1985 to 2000, limited exploration and drilling were conducted, primarily focusing on the geology and potential of the West Dome area. In 1995, Llicancabur Mining Ltda. (Llicancabur), a Bolivian mining company, completed a total of 1,001 m of reverse circulation (RC) drilling in nine holes.

In 2000, COMSUR drilled 914.2 m of diamond core drilling in six holes. The results confirmed significant silver mineralization, as reported in earlier drilling programs. Noteworthy intersections include 52 m grading 103.4 g/t Ag from a depth of 24 m in hole DDH-1 and 30 m grading 62.9 g/t Ag from a depth of 8 m in hole DDH-5.

In 2020, a Bolivian private company, Granville, was granted the PELs in the Carangas area.

In April 2021, NPM announced an Mining Association Contract (MAC) with Granville to jointly explore and develop the Carangas Property. According to the agreement, NPM will cover 100% of future expenditures for exploration, mining, development, and production activities and will receive 98% of operating profits once the Project moves to mining production.

6.2 Property Exploration History

Exploration of the Carangas Project has occurred since the mid-1500s, building on a long history of colonial silver mining in the area. The modern exploration program began with reconnaissance mapping and sampling of historical mine dumps in 2019, followed by progressively more detailed surface and underground mapping, rock chip sampling, and channel sampling through 2020 and 2021. These campaigns included over 1,000 surface and underground samples collected across both domes and the central valley, confirming widespread silver mineralization across the Property. NPM also mapped and sampled approximately 2.4 km of accessible historical underground workings. From 2021 to 2023, exploration expanded to include large-scale geophysical surveys, ground magnetometry, and three-dimensional (3D) Bipole-Dipole induced polarization-magnetotellurics (IP-MT), which identified strong chargeability and magnetic anomalies consistent with known mineralized zones and highlighted several untested targets for future drilling. This sequence of modern exploration work provides the basis for the current understanding of the Carangas mineral system.

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6.3 Historical Resource Estimates
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The Carangas Project had an inaugural Mineral Resources in 2023, which served as the basis for the 2024 PEA Technical Report. Since then, no additional drilling or geological updates have been carried out in the project area.

For the current PEA update, NSR and AgEq were updated considering the new metal prices and technical assumptions. A new Resource pit was calculated that provides the basis for the updated 2026 Mineral Resource estimates. All other reported Mineral Resources are superseded by the estimates stated in this report.

6.4 Production

Mining in the district is believed to have commenced in the sixteenth century and continued intermittently until the early twentieth century, however, there was not record of production data. The Project, particularly the West Dome area, contains historical surface workings, underground mine adits, shafts, and associated processing and smelting infrastructure. Currently, there is no active mining.

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7 Geological Setting and Mineralization
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7.1 Regional Geology
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The Bolivian Central Andes hosts a variety of mineral deposits grouped into distinct metallogenic belts (Figure 7-1). This region hosts epithermal deposits of silver, gold, lead, zinc, and copper. From the Western Cordillera to the east, into the Altiplano, several small copper deposits are hosted in red sandstones of Tertiary age, and some gold deposits associated with subvolcanic intrusions of Tertiary age are hosted in Paleozoic metasediments.

The Eastern Cordillera (or Cordillera Oriental) hosts numerous tin-silver-lead-zinc-tungsten-gold-antimony-bismuth deposits, traditionally known as the Bolivian Silver-Tin Belt, which stretches more than 900 km in length from Peru in the north, through Bolivia, to Argentina in the south, trending from northwest to north-south. The Bolivian Silver-Tin Belt is a significant metallogenic belt hosting numerous super-large silver-tin deposits, including Cerro Rico, Silver Sand, Llallagua, Huanuni, Pulacayo, and Potosí.

Mineral deposits in the Bolivian Central Andes are genetically related to Miocene and Pliocene subvolcanic intrusions of dacitic-rhyolitic composition. Mineralization occurs as veins, veinlets, stockworks, and dissemination hosted in Paleozoic and Mesozoic sedimentary rocks, Cenozoic volcanic rocks, and Paleozoic to Mesozoic plutons.

The Carangas Project is located within a region shaped by long-term subduction of the Nazca Plate beneath the South American Plate. This tectonic setting led to widespread Tertiary volcanic and intrusive activity across the Altiplano, including caldera complexes, volcanic domes, ignimbrites, and subvolcanic intrusions. These magmatic events provided the heat and fluids required to form large hydrothermal systems, which are responsible for the numerous silver, gold, and base-metal deposits found throughout the region.

In the Carangas region a close spatial and genetic relationship between epithermal silver-gold mineralization and deeper porphyry-style systems are observed. Vertical metal zoning is a characteristic feature, with silver-rich mineralization typically occurring at shallow levels, transitioning downward into gold-dominant epithermal mineralization and, in some cases, copper-bearing porphyry systems at greater depth. Structural corridors, caldera margins, and dome complexes play an important role in focusing fluid flow and controlling mineral deposition.

The Carangas Project fits well within this regional framework, displaying extensive epithermal mineralization associated with volcanic centres and showing evidence of vertical zonation and system longevity, both of which support continued exploration potential at depth and along strike.

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Figure 7-1:          RegionalGeology Plan

Source: NPM, 2023

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7.2 Project Geology
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The Central Andes hosts a higher density of volcanoes of ages from Tertiary to Quaternary than any other area in the world. The Carangas region hosts a caldera system of Tertiary age, which is formed over a basement consisting of a moderately deformed Triassic-Lower Jurassic crystalline bedrock and evolved from the Upper Oligocene to Lower Miocene period, as indicated by radiometric dating (Ponce and Avila, 1965). The proposed Carangas caldera is a circular structure with a central dome (Interior Caldera) approximately 20 km in diameter, surrounded by rings of lava domes 25 km to 30 km from the centre (Exterior Caldera) (Figure 7-2). The central dome is interpreted as a resurgent volcanic centre with mineralization in the ring zone, including mineralization systems in Carangas, Negrillos, and Todos Santos.

The Carangas deposit is located at the southwest corner of the Carangas basin, a caldera in the Carangas Exterior Caldera system. Geomorphologically, it consists of two prominent hills, the West Dome and the East Dome, and a valley between them, the Central Valley (Figure 7-3). The two domes are more than 100 m above the surrounding fluvial plains. Near the south end of the Central Valley, a small, outcropped hill is known as the South Dome. Historically, the West Dome is referred to as Espiritu Santo Hill, while the East Dome is referred to as San Antonio Hill.

Based on the results of detailed surface geological mapping and logging of drill cores by the NPM project geologists, the Carangas deposit is interpreted as an epithermal silver-gold mineralization system centred by a rhyolitic diatreme filled with magmatic breccia in the shape of an inverted conical structure spanning from the top of the West Dome eastwards towards the Central Valley (Figure 7-4). The diatreme cuts through the older country rock of volcanoclastic rocks or lithic tuffs of dacitic composition in the upper part and andesitic composition in the lower part of the Carangas Formation.

The upper part of the diatreme is exposed on the top of the West Dome, and three types of rock were identified: hydrothermal breccia, heterolithic breccia, and sandy tuff. To the west of the diatreme breccias, spotty outcrops of dacitic to rhyolitic dykes with flow banding textures are exposed on the surface. The rhyolite dykes roughly strike in a north-northwest direction, and the flow bandings generally dip west at high angles.

The Central Valley is fully covered by young fluvial sediments, ranging in thickness from a few meters to 50 m. Logging of drill cores indicates the rock types beneath the valley are mainly altered phreatic breccia and lithic tuff. To the south, at the South Dome, the outcropped rocks are mainly altered phreatic breccia. Rocks in the East Dome are mainly altered lithic tuffs, likely the overlying phreatic breccias of diatreme already eroded.

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Figure 7-2:          ProjectGeology Plan

Source: NPM, 2023

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Figure 7-3:          CarangasLocal Geology

Source: NPM, 2023

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Figure 7-4:          CarangasDiatreme Structure Section View – Looking NW

Source: NPM, 2023

7.3 Deposit Descriptions

The mineralization of Carangas consists of a diverse suite of metallic sulfide minerals and gangue minerals, occurring as veins/veinlets, breccia fillings, and dissemination. The Company had a joint research program with the Universidad Mayor de San Andrés (Major University of San Andres, UMSA), in La Paz to study the mineralization style and alteration of the Carangas deposit. At least three hydrothermal mineralization phases and one supergene event are identified in the project area.

Mineralization is controlled by the temperature and pressure of the hydrothermal system, i.e., the depth below ground surface or the distance from the heat source generated by rhyolitic intrusions. Three zones of mineralization can be recognized, each containing different metals. The Upper Silver Zone is near the surface and is dominated by silver, with moderate amounts of lead and zinc. Below the Upper Silver Zone, the Middle Zinc Zone is dominated by zinc plus minor silver and lead. The Lower Gold Zone is dominated by gold, with small amounts of silver, copper, and zinc (Figure 7-5). The three mineralized zones are summarized in Table 7-1 and described below.

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7.3.1 Upper Silver Zone
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The Upper Silver Zone formed under relatively low-temperature and low-pressure conditions, at a depth of approximately 150–200 m below the surface. The zone has dimensions of approximately 1,000 m in the east-west direction by 800 m in the north-south direction and extends across the entire area of West Dome, Central Valley, East Dome, and South Dome of the Carangas deposit. It is interpreted as the distal phase of a hydrothermal alteration and mineralization system arising from rhyolitic intrusions at depth in the Central Valley area.

A mineralized horizon up to 50 m thick occurs at the top area of the West Dome. This horizon comprises hydrothermal breccia containing altered rhyolite clasts cemented by low-temperature chalcedonic silica; heterolithic breccia comprising clasts of various lithologies in a matrix of fine debris similar in lithology to the clasts; and unlithified, loose sandy tuff layers and lenses with sedimentary beddings. These three rock types are intercalated with one another. The hydrothermal breccia generally contains a higher grade of silver compared to heterolithic breccia and sandy tuff. When the cementing chalcedony in a hydrothermal breccia appears grey or dark, it may contain up to 1,000 ppm Ag. Due to erosion, the current thickness of this silver-lead horizon is from a few meters up to 50 m thick.

7.3.2 Middle Zinc Zone

As the temperature and pressure of the hydrothermal system increase with depth below the Upper Silver Zone, the grades of silver and lead in the mineralization decrease. In contrast, zinc grades rise with low grades of copper and gold locally in the lower portion of the zone.

The dissemination of marmatite and veining of honey sphalerite, galena, chalcopyrite, pyrite, siderite, and a small amount of silver sulfosalts characterizes mineralization in the Middle Zinc Zone. This zinc-dominated zone is generally from 150 m below surface with a thickness of tens of meters up to 150 m. The Middle Zinc Zone is interpreted as the peripheral zone close to the core Lower Gold Zone, formed in a higher temperature/pressure environment in the vicinity of rhyolitic intrusions.

7.3.3 Lower Zone
7.3.3.1 Lower Gold Zone
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The Lower Gold Zone lies below the Middle Zinc Zone. Mineralization in this zone is characterized by the dissemination of pyrite and sulfides veining of pyrite and chalcopyrite plus a small amount of galena and sphalerite hosted in strongly argillic-sericitic altered phreatic breccia and rhyolite intrusions. This gold zone generally begins at a depth of 200 m and extends to depths greater than 800 m. It has a lateral extent of up to 400 m and is largely confined to the diatreme pipe, with partial lateral extension into the surrounding older volcanoclastic rocks. Laboratory studies indicate that gold occurs mainly in the form of free electrum, in minor amounts as native gold, and very sparsely as iron (gold) sulfides, gold-silver sulfides and galena (gold). The grade of gold generally increases with depth and is highest around an elevation of 3500 m in the middle part of the gold zone. The gold grade declines with increasing depth, however, the copper grade is relatively higher than in the upper portion. The higher temperatures and pressures of hydrothermal activity at depth likely induce this zoning of metals.

Gold mineralization is fully controlled by the diatreme pipe structure, which is associated with rhyolitic dyke intrusions and perfectly overlays with the induced polarization (IP) chargeability anomaly in the Central Valley area. This coincidence may imply that other IP chargeability anomalies beyond the drilled area could represent additional mineral potential and warrant future drill testing.

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7.3.3.2 Lower Copper Zone
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This zone is located below all other structures and is not economically viable. The Lower Copper Zone was not considered for the Mineral Resource Estimate.

Table 7-1          Summaryof Carangas Mineralized Zones

Zone Size* Style<br> of mineralization Mineralization
Upper<br> Silver Zone Approximately<br> 200 m from surface. Dimensions:<br><br> 1000 m L × 800 m W × 200 m T Disseminated<br> silver+lead suphides in the matrices of breccia in the top portion of diatreme (surface of West Dome) and veining plus stockworks<br> of silver+lead+zinc sulfides hosted in diatreme breccia and older volcanoclastic rocks. Silver<br> (lead, zinc)
Middle<br> Zinc Zone 700<br> m L × 600 m W × 150 m T Disseminated<br> sphalerite and veining of zinc plus a minor amount of silver and lead sulfides hosted in the diatreme breccia and in the surrounding<br> older volcaniclastic rocks Zinc<br> (lead, silver)
Lower<br> Gold Zone 400<br> m L × 400 m W × 600 m T Veining<br> of copper-silver-zinc sulfides and disseminated pyrite hosted in diatreme breccia and rhyolitic intrusions, as well as surrounding<br> older volcaniclastic rocks. Gold<br><br> (copper±silver-zinc)

Note: L – length; W – width; T – thickness. Source: NPM, 2023

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Figure 7-5:          MineralizedZones by Metal Zoning- Oblique Section

Source: SLR, 2023

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8 Deposit Types
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8.1 Deposit Model
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The Carangas Project is interpreted to be part of a large epithermal precious-metal system developed within the Central Andean volcanic arc of western Bolivia, a region well known for hosting silver- and gold-rich epithermal deposits. Mineralization at Carangas is associated with Tertiary volcanic rocks, including volcanic domes, breccias, and related intrusive phases, and shows a clear vertical metal zoning with silver-dominant mineralization near surface and gold-dominant mineralization at depth. This vertical zonation, together with widespread hydrothermal brecciation, strong alteration, and a close spatial relationship to volcanic centers, is consistent with a low- to intermediate-sulfidation epithermal deposit model. Regional studies of the Andean Altiplano indicate that such systems commonly form above deep magmatic heat sources, with mineralization focused along structural corridors and caldera-related features. At Carangas, geophysical data, drilling, and surface exposure suggest a vertically extensive hydrothermal system that remains open laterally and at depth, consistent with epithermal systems elsewhere in the Central Andes (e.g., USGS Andean epithermal models; Sillitoe, 2010).

Carangas is located within the Cordillera Occidental belt, close to its eastern limit with the Andes Altiplano. The Cordillera Occidental of Bolivia, along with the Altiplano and Cordillera Oriental, altogether known as the Central Andes, is part of the Andean Cordillera, a convergent plate margin (USGS and GEOBOL, 1975). The Cordillera Occidental is defined by a chain of late Miocene to recent volcanic peaks stretching more than 750 km in length and some 40 km wide (Arce, 2009) that straddles the Bolivia-Chile border. This volcanic arc and associated granitic plutonic rocks of the Coastal Batholith in northern Chile and southern Peru (USGS and GEOBOL, 1975) were emplaced in and cut a Jurassic-Cretaceous aged eugeoclinal-miogenclinal mélange of volcanic flows. Ash flows with associated sedimentary rocks (sandstone, siltstone, conglomerates, tuffaceous sediments and tuffs) all developed over Paleozoic-aged basement rocks (Figure 8-1) (Arce, 2009).

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Figure 8-1:          DepositModel

Source: Arce, 2009

8.2 Deposit Mineralization

The project area mineralization is a silver-gold polymetallic epithermal deposit of low-intermediate sulfidation associated with a rhyolitic maar diatreme cutting into volcanic and volcaniclastic country rocks of Oligocene to Miocene age. The upper portion of the Carangas deposit represents a low sulfidation zone, characterized by argillic and propylitic alterations as well as mineralization of sulfide minerals of silver, lead, and zinc, featured by argentiferous galena, silver sulfosalts, minor native silver, galena, sphalerite, and various gangue minerals, including crustiform-coloform chalcedony, banded chalcedony, smectite, zeolites, carbonates, and chlorite.

To depth, the low sulfidation zone gradually transitions into an intermediate sulfidation zone at a depth of approximately 200 m with sericitic and phyllic alteration and mineralization dominated by gold and a small amount of copper, represented by minerals of electrum, chalcopyrite, pyrite and native gold. Recent microscopic studies conducted in 2022 have further identified the presence of other copper minerals, including enargite (Cu3AsS4) and famatinite (Cu3SbS4). The zone of intermediate sulfidation extends from a depth of approximately 200 m to a depth of more than 700 m. The mineralization in this zone is mainly controlled by the diatreme structure and the intrusion of rhyolite.

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9 Exploration
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9.1 Sampling and Mapping
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The silver mining history of Carangas dates back to the 16th century of the colonial era, evidenced by the widespread historical mining workings and dumps. Systematic exploration programs were completed at Carangas since 2019, including surface and underground (UG) geological mapping, rock chip sampling, and mine dumps sampling, geophysical ground magnetometry surveying, and IP surveying. These exploration programs are summarized in Table 9-1.

At Carangas, 1,076 samples have been collected from an area covering 2 km^2^, including chip and mine dump samples. The sample campaign covers the entire Carangas deposit. Anomalous silver results occur in zones on both domes, confirming extensive silver mineralization at Carangas.

Table 9-1          Summaryof Exploration Programs at Carangas

Year Type of Work Conducted by Description Number of Samples Collected
2019 Surface<br> and UG mapping NPM Grab<br> mine dump samples 268
2020 Surface<br> and UG Mapping NPM Grab<br> dump, surface chip and underground channel sampling 729
2021 Surface<br> and UG Mapping NPM Underground<br> channel sampling 79
2021 Ground<br> magnetometry survey Arce<br> Geofísicos 309.8-line<br> km, 67 N-S lines, 100 m line spacing n/a
2022 3D<br> Bipole-Dipole IP-MT survey Southern<br> Rock Geophysics S.A. 149.2-line<br> km, approx. 10 km^2^ n/a
2022-2023 3D<br> Bipole-Dipole IP-MT survey Southern<br> Rock Geophysics S.A. 28.6-line<br> km, approx. 29 km^2^ n/a
2022 UG<br> mapping NPM n/a

Source: NPM, 2023

During the due diligence study of the Property in 2019, reconnaissance geology mapping and an intensive sampling program of historical mine dumps were carried out to provide an initial understanding of the geology and mineralization of the deposit. Subject to the size of the mine dump, grab samples were randomly selected and collected at 10 m spacings. At least one grab sample was collected if the mine dump was less than 10 m in diameter. The weight of each sample is between 2 kg and 4 kg. A total of 268 samples were collected, of which 233 (86.94%) returned assay results between 30 g/t and 1,950 g/t Ag with an average grade of 270 g/t Ag.

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During detailed surface geology mapping in 2020, the Project collected 383 rock chip samples from 55 outcrops. Chip samples were collected continuously at 2 m intervals of up to 5 cm deep and 10 cm wide along sample lines oriented approximately perpendicular to the strike direction of mineralized structures, for a total length of 769 m. Of the 383 chip samples collected, 117 returned grades between 30 g/t and 2,350 g/t Ag, with an average grade of 160 g/t.

Most of the historical underground mining workings are located at the West Dome, and the Project surveyed and mapped all accessible historical mining adits, totalling 2.4 km in six underground adits. Chip samples were taken continuously at 2 m intervals along the walls of underground workings, and each weighed approximately 2-5 kg. A total of 425 samples were collected, of which 112 samples (26.35%) returned assay results between 30 g/t and 1,060 g/t Ag with an average grade of 122 g/t Ag.

During the site visit, the QP examined the mineralized outcrops and sampling sites at Carangas and concluded that the sampling is representative of the mineralization. The QP did not identify any factors that could have caused sample bias.

9.2 Geophysics

The Project completed geophysical surveying programs, including ground magnetometry and Offset (3D) Bipole-Dipole IP-MT surveying at Carangas in 2021, 2022, and 2023.

The ground magnetometry surveying was conducted by Arce Geofísicos, based in Lima, Peru, from November 2021 until January 2022. A total of 309.8-line km in 67 north-south lines (100 m spaced lines) were completed to cover the entire Carangas area.

Results of the magnetometry survey at Carangas show a prominent low magnetic response approximately centred in the area of the Carangas village, as well as areas with surface mineralization and alteration exposed at West Dome, Central Valley, and East Dome. Magnetic inversion modelling indicates that the lower magnetic response extends to depths of more than 1,000 m. The vast magnetic low beyond and to the north of the drilled area, which includes the West Dome, Central Valley, and East Dome, may indicate the potential for additional mineralization and could justify drill testing in future drilling campaigns.

The 3D Bipole-Dipole IP-MT Surveying at Carangas was conducted in two separate stages:

The first stage was a pilot test carried out in the period July-September 2022 by Southern Rock Geophysics S.A., based in Santiago, Chile, centred on the drilled area, covering approximately 10 km^2^, aiming to understand the geophysical signature of the known mineralization.

The results of the test surveying are very coherent in that multiple chargeability anomalies were identified in the surveyed area, with the strongest anomaly perfectly overlaying the known gold mineralization in the Central Valley (Figure 9-1), which may imply that IP surveying is an effective method for identifying targets of alteration and sulfidation at Carangas. Other chargeability anomalies beyond the Central Valley could be additional mineralization systems.

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The same contractor conducted an expanded second-stage surveying from October 2022 to January 2023 to cover the entire Carangas caldera basin, totalling 130,993-line m over 29 km^2^, aiming to identify additional chargeability anomaly targets.

The expanded surveying confirmed the anomalies in the central drilled area of Carangas and identified additional anomalies across the Carangas caldera basin. The most prominent anomaly of chargeability lies to the north of West Dome, trending roughly NNW, parallel to the strike of the surface mineralized structures, with the intensity of the anomaly increasing from a depth of 200 m below surface (elevation 3,700 masl), similar to the geophysical response of the gold mineralization in the Central Valley area. This anomaly is a good target for future drilling campaigns.

Figure 9-1:          IPChargeability Anomalies of the Carangas Area

Source: NPM, 2023

9.3 Exploration Potential

Based on the Mineral Resource estimate outcomes, the QP recommends that additional drilling be undertaken. There is good potential to expand the Mineral Resource base and increase confidence in the Mineral Resource, which is required to make informed investment decisions.

The QP notes that at this stage, any of the related exploration potential mentioned in this report has not been reported as Mineral Resources, and there is no guarantee that Mineral Resources will be defined through further exploration.

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10 Drilling
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10.1 Introduction
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This section details all drilling activities completed on the Project and all the data provided to the SLR QP to estimate the Mineral Resource. Since 2021, 189 boreholes were drilled at the Carangas Project, totalling 81,145 m of diamond core drilling. These drill holes were used to compile the Mineral Resource. Drill hole spacing averages 50 m by 50 m in the most densely drilled areas and increases to 100 m by 100 m on the peripheries of the deposit and at depth. A summary of drilling data within the Carangas Mineral Resource area is presented in Table 10-1, and hole locations are shown in Figure 10-1.

The drilling were designed to test surface mineralization at the West and East Domes, delineate covered mineralization in the Central Valley, and define the extent of silver- and gold-bearing zones on a nominal 50 m × 50 m grid in the core of the deposit, with wider spacing toward the margins and at depth.

Drilling successfully intersected broad zones of mineralization, confirming a large epithermal system with silver-dominant mineralization near surface and a deeper gold-dominant zone. The results demonstrate good lateral continuity of mineralization within the drilled area and provide sufficient data to support the current Mineral Resource estimate. Overall, the drilling results (Figure 10-1) are considered reliable and appropriate for a PEA-level evaluation, while additional infill, step-out, and oriented drilling will be required to improve confidence in mineralization geometry and support future resource upgrades. Table 10-210-2 Drilling Intercepts examples presents some drilling intercepts used in modelling and Mineral Resource estimates.

Table 10-1:          CarangasDrilling History

Year Drilling Phase Carangas
Holes Meters
2021 Phase<br> 1 - Discovery Drilling 13 3,790.4
2021 Phase<br> 2 - Discovery Drilling 22 9,420
2022 Phase<br> 3 - Resource Definition Drilling 115 50,310.92
2023 Phase<br> 4 - Resource Definition Drilling 39 17,623.5
Total 189 81,145

Source: NPM, 2023

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Figure 10-1:          DrillholeCollar Location Plan

Source: NPM, 2023

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Figure 10-2:         DrillholeCross-section

Source: SLR, 2026

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Table 10-2:            DrillingIntercepts Examples

Hole number From To Ag (g/t) Au (g/t) Cu (%) Pb (%) Zn (%)
DCAr0034 158.0 159.3 191.00 0.01 0.03 0.58 1.23
DCAr0034 160.6 161.9 397.00 0.01 0.03 0.61 1.49
DCAr0034 165.2 166.3 46.40 0.04 0.01 0.67 1.49
DCAr0034 166.3 167.3 49.20 0.04 0.01 0.58 1.20
DCAr0034 167.3 168.4 51.00 0.01 0.01 0.29 0.83
DCAr0034 168.4 169.6 32.90 0.07 0.01 0.25 0.89
DCAr0034 527.6 528.6 149.00 0.55 1.20 2.29 6.53
DCAr0034 529.0 530.3 51.30 0.48 0.47 1.34 2.89
DCAr0034 530.3 531.4 12.50 0.13 0.01 0.78 2.04
DCAr0034 531.4 532.5 11.90 0.08 0.03 0.76 2.13
DCAr0034 532.5 533.6 13.25 0.05 0.03 1.13 2.87
DCAr0094 47.7 48.9 12.40 0.01 0.00 0.32 0.68
DCAr0094 48.9 50.1 16.95 0.01 0.00 0.36 0.99
DCAr0094 50.1 51.4 16.10 0.01 0.00 0.36 0.99
DCAr0094 54.0 55.2 18.65 0.01 0.00 0.29 0.74
DCAr0094 246.2 247.4 20.30 0.31 0.05 1.25 0.71
DCAr0094 251.2 252.5 15.20 0.08 0.06 0.71 0.75
DCAr0094 256.0 257.2 10.70 0.11 0.06 0.54 2.16
DCAr0094 263.4 264.5 76.70 0.25 1.33 0.58 2.87
DCAr0094 303.8 305.0 12.25 0.16 0.14 0.44 0.92
DCAr0162 295.4 296.6 11.00 0.16 0.01 1.11 1.66
DCAr0162 297.8 299.0 16.00 0.34 0.02 2.17 1.28
DCAr0162 299.0 300.2 11.00 0.13 0.06 0.88 0.85
DCAr0162 301.7 303.0 13.00 0.02 0.09 0.43 1.14
DCAr0162 368.0 369.2 27.00 0.19 0.02 0.37 0.65

Source: SLR, 2026

10.1.1 Drill Methods

Drilling was completed using a conventional wire-line diamond drilling technique with a triple tube core barrel inside the drilling rods to produce HQ, NQ, or PQ size diamond core. Each drill run was 3 m in length. The drill core was placed in plastic/wood core trays (each holding approximately 4 m of drill core) after extraction from the core barrel, where each run was marked and labelled.

10.1.2 Drilling Locational Data

Technical surveyors surveyed all drill hole collar locations using the Real Time Kinematic (RTK) Global Positional System (GPS) method. The QP noted that all drill collars align well with the topography. The drilling teams utilized a Reflex EZ-track™, SPT GyroMaster™, and SPT Core Retriever™ instruments to measure deviations in azimuth and inclination angles. The measurements were taken approximately every 30 m along the drill trace.

10.1.3 Logging Procedures

The NPM developed logging and sampling procedures based on the experience of the technical team and industry standards. Geological logging included lithology, alteration, weathering, structure, and mineralogy. During the QP site visit, several representative intervals were checked to assess logging quality. No issues were noted. Core loss was noted as problematic in overburden/saprolitic zones and in voids (due to historical artisanal mining activities). The overburden zone is considered a waste zone and is not considered in the Mineral Resource estimate.

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Core photography and core recovery measurements were carried out by assistants under a geologist’s supervision and digitally recorded into the MX Deposit system. During the site visit, the QP reviewed recent core photos and noted that the photo quality aligns with industry expectations.

The core is stored in two different core yards at the project site. The core samples, pulps, and coarse rejects are properly stored at the core shack at the project site.

The QP noted that the technical team maintains a well-organized workflow and a good core storage plan on-site. A new core yard is being prepared to store the core for future drilling programs.

10.1.4 Recovery

Core recoveries were calculated by measuring the length of the core recovered from each three-meter run. The average core recovery is higher than 95%,and more than 89% of core intervals have a core recovery of 95% or greater. Recovery was low in the voids (historical mining activities) and overburden (mining dumps and fluvial sediments). In the opinion of the QP, there are no known factors regarding drilling, sampling, and core recovery that could materially impact the accuracy and reliability of the results.

10.1.5 Sample Length/True Thickness

Drilling at the Carangas Project has outlined a large epithermal system with both silver-rich zones near surface and gold-dominant mineralization at depth; however, the detailed geometry and structural controls of these mineralized bodies remain only partially defined in the current geological model. Because the mineralization occurs in a combination of volcanic units, breccias, and dome-related intrusive rocks, and because most drill holes were not oriented, the true dip, strike, and continuity of the mineralized zones cannot yet be confidently interpreted from the available data. The current sample length varies from 1 m to 1.5 m.

As a result, the relationship between the drill hole sample lengths and the true thickness of the mineralized intervals cannot be reliably determined at this stage. All intercepts reported to date represent down-hole lengths, which may not reflect the actual thickness of the mineralized bodies. Establishing true thickness will require additional drilling, including oriented core and improved structural logging, together with refinement of the 3D geological and mineralization models.

10.2 Comments on Drilling

During the site visit, the QP discussed the diamond drilling, core handling, and drilling techniques and considered them appropriate and consistent with the Mineral Resource estimation and classification. Further information is provided in Section 12.

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11 Sample Preparation, Analyses, and Security
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11.1 Introduction
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The details of the sample preparation, analytical methodology, and sample security protocols in place for core samples from the exploration programs carried out by NPM are included in this section.

NPM managed the 2021-2023 drill programs of the Carangas Project following the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Mineral Exploration Best Practice Guidelines (CIM, 2019) and internal working protocols. All drill cores at the Project were geologically logged and sampled by the NPM’s exploration team at its core processing facilities in accordance with the NPM’s core logging and sampling protocols. A total of 58,215 half-sawn core samples were taken and submitted for preparation and analysis.

In QP’s opinion, the QA/QC program, as designed and implemented by NPM, is adequate, and the assay results within the database are suitable for use in a Mineral Resource estimate.

11.2 Sample Collection

NPM systematically sampled and analyzed all drill core from the 2021-2023 exploration drill programs. Core sampling totalled 58,215 half-swan core regular samples submitted for preparation and analysis. Contracted diamond drillers used HQ-size coring equipment for 189 drill holes and various-sized tubes (TS-HQ-NQ) for deep drill holes exceeding 500 m in length.

All drill holes were geologically logged and sampled by technical field personnel at the Carangas facility in accordance with core logging and sampling protocols. Geological logging includes the detailed recording of lithology, alteration, mineralization, structure and RQD measurements. Rock codes were developed to increase the quantity and quality of geological data and to build a robust geological model for the deposit. Logging data was entered directly into the MX Deposit software developed by Seequent. MX Deposit is an industry-standard software suite that integrates a drill core logging module, a drill hole database, and a QA/QC tool for real-time monitoring of analytical result quality.

The driller contractor's staff transported core boxes from the drilling site to the shed. The core was cleaned or washed, core blocks were checked, and meter marking was completed. Samples were generally one meter in length from one whole depth meter to the next, except for where a lithological contact or alteration change was noted. Samples were a maximum of 1.5 m and a minimum of 1.0 m (Figure 11-1). Geologists also mark noticeable geological, structural, and alteration contacts and intervals of poor core recovery (voids and core loss). The core was photographed wet, using a camera mounted in a frame to maintain a consistent angle and distance from the core.

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Figure 11-1:         DrillCore Box Example – Drill Hole DCAr0171

Source: NPM, 2023

On completion of logging and sample selection, all core boxes were transported to the core saw shed. The core was cut using a diamond saw, and unconsolidated material was split using spoons or trowels. Each sample interval was placed in a plastic bag with a sample ticket. Sample intervals are cross-checked with the sample tag book and the pre-labelled sample bag (Figure 11-2). The outer portion of the tear-off sample tag is affixed to the core box at the start of the sample interval, and the inner tear-off tag is placed in the sample bag.

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Figure 11-2:         CoreCutting and Sample Bag

Source: NPM, 2023

Once sampling is complete, geologists check the samples and seal the plastic sample bags with staples and tape. QA/QC samples are inserted into the sample sequence in accordance with the Project’s QA/QC protocols. Then, every 8 to 12 sample bags are placed into a large poly-weave sample bag for shipping to the laboratory for preparation.

During the site visit, the QP observed that NPM employees understood core sample preparation procedures well. The QP also observed that all equipment used for core sample preparation was of reasonable quality and in line with industry standards.

11.3 Assay Laboratory Sample Preparation and Analysis

All drill core samples collected by NPM between 2021 and 2023 were dispatched to ALS in Oruro, Bolivia, for sample preparation and then to ALS in Lima, Peru, for geochemical analysis. ALS Oruro and ALS Lima are part of ALS Global, an independent commercial laboratory specializing in analytical geochemistry services. Both laboratories are certified in accordance with the International Organization for Standardization (ISO) and International Electrotechnical Commission (IEC) “General requirements for the competence of testing and calibration laboratories” (ISO/IEC 17025:2017).

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All samples are prepared in accordance with ALS preparation code PREP-31 and follow the main standard procedure as outlined below:

· Samples<br> were dried and crushed to 70% less than 2 m.
· A<br> 250 g riffle split was taken and pulverized to greater than 85% passing a 75 µm sieve<br> prior to aliquot selection for digestion and analysis.
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· The<br> pulp samples are transferred to ALS Lima for geochemical analysis.
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· Samples<br> were submitted for trace level 51 elements analysis comprising aqua regia digest with Inductively<br> Coupled Plasma-Mass Spectroscopy (ICP-MS) finish, ALS Code ME-MS41. Over-limit samples returning<br> results Ag >100 ppm, Pb >10,000 ppm, Cu > 10,000 ppm, and Zn >10,000<br> ppm were sent for grade analysis by aqua regia digestion with Inductively Coupled Plasma-Atomic<br> Emission Spectroscopy (ICP-AES) finish, analysis ALS code OG46. Samples returning Ag assay<br> results greater than 1,500 g/t were analyzed by fire assay and gravimetric finish using the<br> ALS code Ag-GRA21. Samples returning values over 10,000 ppm Ag were analyzed by high-precision<br> analysis through fire assay and gravimetric finish, ALS code Ag-CON01. Gold by fire assay<br> and atomic absorption spectrometry (AAS) analysis, ALS Code Au-AA25, was performed on drill<br> core samples selected from long drill holes exceeding 500 m in length.
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For the 2023 drill program, no trace-level multi-element ICP analysis of drill core samples was performed in order to shorten turnaround time and save costs.

11.4 Bulk Density

Specific gravity (SG) measurements are completed by technical personnel as part of routine core processing procedures. A total of 5,367 measurements were completed with a mean SG of 2.19 for core intervals selected across various lithologies and alteration types in both mineralized and non-mineralized drill cores at a rate of 8% to 9% of the total core samples. Measurements are carried out at a dedicated density weighing station using Archimedes' principle, whereby water displacement is used to calculate approximate volume (Figure 11-3). To prevent water absorption by porous drill cores, the core interval is waxed prior to immersion in water. Weighing scale calibration is performed daily before each measurement.

The bulk density of a sample is calculated by multiplying the SG by the density of water (1 g/cm^3^).

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Figure 11-3:         SpecificGravity Measurement

Source: NPM, 2023

11.5 Quality Control Data

NPM has established comprehensive QA/QC procedures and protocols covering sampling, preparation, and geochemical analysis. All drilling programs completed on the Project were performed with mandatory insertions of CRMs, blanks, and duplicates into normal sample sequences on a batch-by-batch basis. NPM monitors silver, gold, lead, zinc, and copper assay values in CRMs, blanks, and duplicates.

NPM provided QA/QC data for drilling completed during the 2021 – 2023 exploration drilling campaigns. The QA/QC samples comprise 24% of all Carangas samples submitted to the laboratory. The QP is of the opinion that adequate QA/QC protocols were in place for the entire drilling used to compile the Mineral Resource estimate.

The QA/QC procedures utilized various control samples, including CRMs; coarse and pulp blanks; coarse, field (1/4 core) and pulp duplicate samples; and umpire pulp duplicate samples. Detailed statistics of QA/QC control samples are presented in Table 11-1.

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Table 11-1:            QA/QCSample Status

Type Number of Samples % of Total Primary Samples
Standards (CRMs) 3,654 6%
Blanks (Coarse and Pulp) 3,038 5%
Duplicates (Coarse, Pulp, and Field) 4,269 7%
Umpire Pulp 3,573 6%
Total 14,534 24%

Source: compiled by SLR, 2023

11.5.1 Certified Reference Materials (CRMs)

In total, six different CRMs have been used for QA/QC testing. CDN-GEO-1901, CDN-ME-1501, and CDN-ME-1603 were discontinued in 2022. All CRMs were supplied by CDN Resource Laboratories of Langley, British Columbia, Canada, with certified values of Ag, Au, Pb, Cu, and Zn. CRM statistics for Ag and Au are presented in Table 11-2.

Table 11-2:            CRMsof the Carangas Project

CRM Ag ppm Au ppm CRMs inserted
Certified value 2SD Certified value 2SD 2021-2023
CDN-GEO-1901 1 0.3 0.036 0.008 408
CDN-ME-1501 34.6 2.3 1.38 0.11 288
CDN-ME-1603 86 3 0.995 0.066 985
CDN-ME-1707 27.9 2.9 2.02 0.214 893
CDN-ME-1902 349 17 5.38 0.42 780
CDN-ME-2003 106 9 1.301 0.135 300

Source: compiled by SLR, 2023

NPM’s internal procedures require that one CRM be inserted for every 20 samples, or at a rate of 5% under a random insertion protocol. CRM performance is monitored on a batch-by-batch basis. A total of 3,654 CRM samples were submitted in 2021 – 2023, equivalent to a rate of 6.0%.

Control charts are used to monitor the analytical performance of individual CRMs over time. CRM assay results are plotted in order of date of analysis. Charts present certified CRM values, an analytical mean line, and control lines for acceptable limits, calculated as the analytical mean of the CRMs plus or minus two standard deviations (2SD). Results outside of three standard deviations are considered failures. These charts show analytical drift, bias, trends, and outside-of-tolerance outliers occurring in the laboratory over time. The analytical mean line shows the variability of the analyzed material.

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Figure 11-4 presents CRM control charts for silver by ICP- MS analytical methods. Sporadic outliers (yellow circles) that are slightly higher or lower than the warning limits do not affect the laboratory procedure's analytical mean, accuracy, and precision. Failed standards were re-assayed and investigated by the laboratory. Comparison between original and re-assayed values proved the accuracy of the laboratory’s original assay results. Overall, the CRMs have a very good performance and support the sample database for the resource estimation process.

Figure 11-4:          ControlChart for CDN-ME-1501 (Ag) (July 2021 – November 2022)

Source: NPM, 2023

Standard CDN-ME-1707 demonstrated relatively poor performance for gold compared with other CRMs. Failed samples were investigated and documented. Occasionally, CRM material may exhibit a concentration of elements that differs from the certified values. It is possible to have precise results that are not accurate. All other CRMs inserted in the same batches passed control limits. CRMs used at Carangas to monitor gold show acceptable analytical accuracy and provide confidence in analytical results for the span of gold grades at the deposit. All major differences were investigated, and appropriate action was taken to fix it and return a robust data control.

11.5.2 Blank Control Samples

Two types of blank material were inserted into the sample sequence prior to delivery to the laboratory. Coarse blanks are used to assess the potential contamination during sample preparation, and pulp blanks are used to assess the potential contamination during geochemical analysis.

The coarse blank material used at the Carangas Project was taken from a quarry located near Oruro. The rock is fresh andesite with porphyritic texture containing quartz, plagioclase, biotite, and hornblende grains. The chemical validation for coarse blanks was developed internally and certified that the material could be used for this purpose. The NPM developed the control limits after reviewing the analytical data, removing outliers, and calculating the analytical mean and standard deviation (SD). The warning limit is set at two standard deviations. The failure limit is set at three standard deviations.

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Overall, 99.5% of the coarse blanks are within the acceptable limits (Figure 11-5). Failed results exceeding three standard deviations were documented and investigated.

Figure 11-5:          ControlChart for Coarse Blank Samples

Source: NPM, 2023

According to QA/QC protocol, pulp blank samples were to be inserted every 50 samples or at a rate of 2%. A total of 1,031 pulp blank samples were inserted from July 2021 to April 2023, representing an overall insertion rate of 2.3%.

Certified pulp blank CRM CDN-GEO-1901 was used between July 2021 and April 2022. Unlike other CRMs, CDN-GEO-1901 did not demonstrate high accuracy (-13.57% difference between analytical mean and certified value) for silver analysis. However, 99% of CDN-GEO-1901 samples were within the control limits (Figure 11-6).

Since April 2022, pulp blanks have been produced from pulverized coarse blank material used to monitor potential contamination during sample preparation.

Figure 11-6:         ControlChart for Pulp Blank CDN-GEO-1901

Source: NPM, 2023

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Overall, 99.5% of pulp blank samples for silver are within two standard deviations of the control limit. It is concluded that there is no systematic contamination during geochemical analysis.

A total of 981 coarse blanks were inserted into the sample sequences for gold fire assays in the period 2021-2023. Only eight coarse blanks returned results above the failure limit of 0.025 ppm Au. Every failure was documented and investigated. A total of 97% of the coarse blank samples analyzed for gold returned with assay results equal to or below 0.01 g/t Au (twice the detection limit of 0.005 ppm Au). No contamination was identified during sample preparation and analysis.

11.5.3 Duplicate Samples

Three types of duplicates were used to monitor the quality of the Carangas drill programs' processes, from sampling through preparation and analysis: twin samples (field duplicates), coarse reject duplicates, and pulp duplicates. A total of 4,269 duplicate samples were taken during the period July 2021 – April 2023. Table 11-3 provides a statistical summary of the Relative Percent Difference (RPD) for the assay pairs between the original and the duplicate of each type of duplicate sample.

Table 11-3:            StatisticalSummary for Duplicate Samples July 2021 – April 2023

Ag ppm
Sample Type Number of samples Corr Coeff < 10% RPD < 20% RPD
Field Duplicate 1,463 0.939 50% 70%
Coarse Duplicate 1,425 0.997 79% 89%
Pulp Duplicate 1,381 0.997 80% 91%
Au ppm
Field Duplicate 683 0.95 55% 66%
Coarse Duplicate 670 0.982 62% 72%
Pulp Duplicate 671 0.984 63% 72%

Source: compiled by SLR, 2023

Field duplicates are generated by the quarter core to monitor the representativeness of the sampling process. The insertion rate is 2% according to the QA/QC protocols, and 1,463 quarter-core duplicates were taken during the 2021-2023 drilling campaigns.

The performance of the field duplicate for silver is presented by the Thompson-Howarth precision plot (Figure 11-7) and the quantile-quantile (Q-Q) plot (Figure 11-8). In both charts, the results are reasonable and support the Mineral Resource database.

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Figure 11-7           PrecisionPlot of Field (1/4 core) Duplicates for Silver Assays

Source: NPM, 2023

Figure 11-8           Quantile-QuantilePlot of Field (1/4 core) Duplicates for Silver Assays

Source: NPM, 2023

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To monitor the sub-sampling or splitting precision during sample preparation, a coarse (reject) duplicate is taken immediately after the first crushing and splitting step. The duplicate reject has a similar weight to the original sample and follows the same preparation process as the original sample. A total of 1,425 coarse duplicates were taken from the 2021-2023 drilling programs, inserted at a ratio of 2%, or one in every 50 samples.

The assay results for silver from the original and duplicate pairs show a high correlation coefficient (R=0.997), and 89% of the sample pairs have an RPD less than 20%, indicating that the sampling and splitting process is highly precise and well representative of the mineralization. In Figure 11-9, the blue dashed lines mark +10% tolerance and red lines +20% tolerance from the black 1:1 line, respectively. The scatter plot shows that nearly all duplicates are within acceptable tolerance.

Figure 11-9           CoarseDuplicate Precision Scatterplot – Silver Assays

Source: NPM, 2023

Both the silver and gold assays from the coarse duplicates demonstrate that the sample preparation process for the Carangas Project is well preserved and acceptable for the Mineral Resource estimate.

Pulp duplicates are used to monitor the precision or repeatability of geochemical analysis. The samples are inserted into the regular sample sequences and are analyzed by ALS (Lima). Pulp duplicates are the second split of the final pulps, with a weight similar to that of the original sample. The required insertion rate is 2%, or one in every 50 samples. A total of 1,381 pulp duplicates were taken for the 2021-2023 drill programs.

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The assay results for silver from the pairs show a high correlation coefficient (R=0.986), with 90% of sample pairs having an RPD of less than 20% and 80% having an RPD of less than 10%, indicating that the laboratory process for geochemical analysis has high repeatability and precision. Figure 11-10 shows the Thompson-Howarth precision plot of pulp duplicates for silver samples in 2021-2023 drill campaigns.

Figure 11-10        PrecisionPlot of Pulp Duplicates for Silver Assays

Source: NPM, 2023

The silver and gold assay results from the pulp duplicates demonstrate high precision (repeatability) in the geochemical analysis by ALS (Lima), and the results are acceptable for database validation.

11.5.4 Umpire Laboratory Samples

To assess the analytical accuracy of ALS (Lima) as the primary laboratory, umpire-check samples were sent to Alfred H Knight Laboratory (AHK) in Lima, Peru, a second accredited laboratory for the check analysis of drill core samples from the Carangas Project during August 2021 – May 2023. AHK is an independent geochemical laboratory certified according to ISO/IEC 17025:2005 and ISO 45001:2018. The required ratio of umpire samples collected from the pulp rejects of normal sample sequences is 5% to 6%, or five to six umpire samples from every 100 primary samples, according to NPM QA/QC protocol. Table 11-4 summarizes the silver and gold assay results of sample pairs of the originals and the umpires for the drill core samples from 2021 to 2023.

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Table 11-4             StatisticalSummary for Umpire Duplicates Samples

Sample Type Element Number of samples Cor Coeff < 10% RPD < 20% RPD
Umpire Pulp Duplicate Ag ppm 2,509 0.986 81% 91%
Au ppm 1,064 0.935 38% 58%

Source: compiled by SLR, 2023

A total of 2,509 umpire samples were assayed for silver by Inductively Coupled Plasma - Optical Emission Spectroscopy (ICP-OES) at AHK, with silver values ranging from 0.2 ppm to 1,725 ppm. The comparison between the original and umpire assay pairs is displayed in Figure 11-11. A correlation coefficient of R=0.986 reveals a strong positive correlation between the original and umpire assay results. A total of 91% of umpire duplicates have an RPD of less than 20%, indicating good reproducibility of silver assay results.

Figure 11-11        UmpirePulp Duplicates Precision Scatterplot for Silver Assays

Source: NPM, 2023

Gold performs significantly worse than silver, likely deriving from free gold in some areas of the Project.

In conclusion, the external laboratory check analysis of silver and gold demonstrates good accuracy and precision of geochemical results produced by ALS (Lima), which supports the database used for resource estimate procedures.

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11.6 Security and Storage
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The technical staff takes custody of drill cores and samples at every step of field exploration and drilling activities. No other people were allowed to enter the working areas and the core storage without pre-approval from the project manager. The Project core is stored in plastic core boxes and transported to the core logging shack. After being logged and sampled, the core boxes are shipped to a secure core yard regularly for permanent storage (Figure 11-12).

Core samples are collected from the drill site at least every 24 hours as part of routine drill site inspections and supervision provided by site geologists. Geological “quick logs”, portable X-ray fluorescence (XRF) analyses, and photographs of each core box are completed during the site inspection and before core boxes are transported to the core logging and sampling facility. Sample bags are transported to ALS (Oruro) by NPM personnel using a truck equipment. The Sample Submission Order is reviewed and signed by ALS staff upon arrival, and the laboratory then assumes custody of the security.

Figure 11-12        SecureCore Yard Storage

Source: SLR QP’s site visit, 2023

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11.7 Comment on Sample Preparation, Analyses, Security and QA/QC
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The QP is of the opinion that the overall QA/QC process is well established and that the results support the Mineral Resources estimation process.

The procedures and protocols employed by the NPM regarding sampling, preparation, sample security, and analysis are in accordance with industry best practices. The QP did not identify any material concerns with the geological and analytical procedures or the quality of the results at the Carangas Project.

The use of different control samples is robust and returns a good variety of verification through the process. The umpire laboratory check analysis shows a high level of reproducibility in the database.

The insertion rate of control samples is 24%, which is higher than the industry benchmark (15% to 20%).

During the site visit, the QP identified that the sample preparation procedures and geology core logging are well established and contributed to a robust database. Good operational procedures are in place for core preservation and storage.

All drill core and samples were handled under controlled conditions, with a continuous chain of custody maintained from the drill site to secure on-site storage and through transport to accredited laboratories, ensuring sample integrity and security of the process. The QP is of the opinion that the results are acceptable and consistent with industry standards and recommends that NPM maintain a continuous QA/QC program for future exploration drill campaigns to maintain the database quality.

11.7.1 Recommendations
· Maintain<br> the current sampling, preparation, analytical, and security procedures for all future drilling<br> and exploration programs, as they follow industry best practices and support reliable Mineral<br> Resource estimation.
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· Continue<br> using accredited, independent laboratories for sample preparation and analysis, and maintain<br> strict chain-of-custody<br> procedures from drill site to laboratory to preserve data integrity and auditability.
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· Retain<br> the comprehensive QA/QC program, including regular insertion of certified reference materials,<br> blanks, duplicates, and umpire laboratory samples, at rates consistent with or exceeding<br> current practice.
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· Continue<br> close monitoring of gold assay performance, particularly in zones where free gold may be<br> present, and promptly investigate and document any QA/QC failures or trends, including re-assaying<br> where necessary.
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· Continue<br> collecting bulk density measurements at representative frequencies across all lithologies<br> and mineralization styles to support future Mineral Resource updates.
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· Ensure<br> that the same level of documentation, supervision, and QA/QC oversight is maintained in all<br> future programs to preserve consistency and reliability of the assay.
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12 Data Verification
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12.1 Geological and Mineral Resource Data Verification
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This data verification discussion herein addresses only the data used to inform the Mineral Resources.

12.1.1 Verification performed by the QP
12.1.1.1 Data Verification Measures
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The QP did not identify any inaccuracies or misrepresentations of the underlying assay results in the database.

The drill database for the Mineral Resource estimate was received in digital format, and the QP conducted a systematic review of the data in Excel and Leapfrog.

The QP conducted a site visit to the Carangas Project in March 2023, viewed outcrops, drill hole locations, historical artisanal mining activities, and core sheds, and held various discussions with the project geologists. The QP examined mineralized drill hole intersections, downhole survey and assay data, acquisition protocols, logging and sample preparation procedures, and quality assurance (QA) procedures (QA) and quality control (QC) results.

NPM supplied the digital topographic file. The 1 m stereo satellite survey was conducted in 2021 by PhotoSat, based in Vancouver, Canada. The QP verified drill collar locations during the site visit and found the Relative Levels (RLs) of the topography at these locations within the expected variations. The QP did not find any inaccuracies related to the topography surface and collar location.

The QP concluded that the data was adequately acquired and validated in accordance with industry best practices.

12.1.1.2 Database Validation

The QP completed systematic data validation steps after receiving the database, including the following checks:

· The<br> collar table was checked for duplicate holes
· Down-hole<br> data (surveys, assays, bulk density, recovery, geology) was verified
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· Hole<br> dips angles and directions
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· Visual<br> inspection of drill hole collars and traces.
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The Carangas drill hole database contains 189 drill holes representing 81,145 m. A total of 58,212 samples were analyzed and comprise the current data for Mineral Resource estimation.

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The QP randomly validated approximately 5% of assay certificates against the database's assay records. The QP did not identify any inconsistencies and opines that the assay database is suitable for geological interpretation and the Resource Estimation process.

12.1.1.3 Validation of Mineralization

During the site visit, the QP viewed outcrops, drill hole locations, and mineralized drill hole intersections. The QP viewed the representative mineralized drill core intercepts listed in Table 12-1 at the core shed located at the Carangas Project site.

Table12-1:            Drill Core Intervals Viewed

BHID FROM TO
DCAr0179 500.00 700.00
DCAr0096 0.00 950.00

Source: compiled by SLR, 2023

The mineralization intervals were verified in drill core intercepts. Figure 12-1 provides typical examples of core mineralization intercepts.

Figure 12-1:         DrillCore Mineralization Intercept Examples

Source: QP’s site visit, 2023

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12.1.1.4 Drill Hole Location Validation
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During the site visit, drill collar locations for DCAr0052, DCAr0156, DCAr00169, and DCAr00171 were checked by handheld GPS and drill hole orientations were checked by compass. Variations of 1m to 3 m were noted, and the SLR QP considers this to be within the expected accuracy of the different measurement systems. Each drill hole was capped and labelled, and easily identified in the field (Figure 12-2).

The QP is of the opinion that drill hole locations and orientation information supplied in the database are of a suitable standard, and the data can be used for Mineral Resource estimation.

Figure 12-2:         Examplesof Drill Hole Collar Field Registration

Source: SLR, 2023

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12.1.1.5 Core Logging, Sampling, and Storage Facilities
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The NPM developed logging and sampling procedures based on the experience of the technical team and industry standards. Geological logging included lithology, alteration, weathering, structure, and mineralogy. During the QP’s site visit, several representative intervals were checked to assess logging quality. No issues were noted. Core loss was noted as problematic in overburden/saprolitic zones and in voids (due to historical artisanal mining activities). The overburden zone is considered a waste zone and is not considered in the Mineral Resource estimate.

Core photography and core recovery measurements were carried out by assistants under a geologist’s supervision and digitally recorded into the MX Deposit system. During the site visit, the QP reviewed recent core photos and noted that the photo quality aligns with industry expectations.

The core is stored in two different core yards at the project site. The core samples, pulps, and coarse rejects are properly stored at the core shack on the project site.

The QP noted that the technical team maintains a well-organized workflow and a good core storage plan on-site.

12.1.1.6 Risks

Key risks associated with data verification are as follows:

· Reliance<br> on historical drilling data (2021-2023) as no new drilling has been completed since the 2023,<br> meaning future changes in geological interpretation will require additional data collection.
· Dependence<br> on contractor-generated topography<br> and surveys, while verified and within tolerance, the Project relies on third-party<br> survey products that require ongoing validation in future updates.
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12.1.1.7 QP Comments on Geological and Mineral Resource Data Verification
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The QP is of the opinion that the drill data is adequate for the purposes of geological interpretation and Mineral Resource estimation within the classifications applied.

The Ausenco QP has reviewed available topography, mine schedule, mineral processing, and climate data to be reasonable and suitable for the purposes of the level of this report. While geotechnical investigations, high-definition topography, and additional processing test work and optimization of the mine plan may be conducted during subsequent stages of the project development, the current available data are considered adequate to support the geotechnical assumptions and technical conclusions presented in this report.

12.2 Metallurgical Data Verification

The QP has reviewed the available metallurgical testwork results and supporting technical data used in this study. The review included an evaluation of laboratory procedures, sample preparation methods, analytical results, and the consistency of metallurgical recovery results with the geological domains and mineralization characteristics of the deposit.

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Based on this review, the QP considers the metallurgical data to be reasonable and suitable for the purposes of this Technical Report. While additional metallurgical testwork and engineering studies may be conducted during subsequent stages of project development, the currently available data are considered adequate to support the metallurgical assumptions and technical conclusions presented in this report.

12.3 Geotechnical Data Verification

The QP has reviewed available topography, mine schedule, mineral processing, and climate data to be reasonable and suitable for the purposes of the level of this report. While geotechnical investigations, high-definition topography, and additional processing test work and optimization of the mine plan may be conducted during subsequent stages of the project development, the current available data are considered adequate to support the geotechnical assumptions and technical conclusions presented in this report.

12.4 Socio-environmental data verification

The QP has reviewed the available socio-environmental studies completed to date to support the Project and considers the data to be reasonable and suitable for the purposes of the level of this report. Identified information gaps can be filled in subsequent stages of the project development and permitting efforts by means of additional studies.

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13 Mineral Processing and Metallurgical Testing
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13.1 Introduction
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A preliminary metallurgical testwork program involving flotation and cyanide leach was started in June 2022 by Bureau Veritas Minerals in Richmond, British Columbia, Canada. Five composite samples were tested. This program was later continued in 2023 by ALS Metallurgy in Kamloops, British Columbia, Canada. During 2023, another three composite samples were collected from the silver zone and gold zone for the comminution testing by ALS Metallurgy in Kamloops. Three reports were issued and some of important results have been described in a NI 43-101 technical report published in August 2023.

· Bureau<br> Veritas Minerals, Metallurgical Testing for Gold, Silver, Lead and Zinc Recovery, New Pacific<br> Metals – Carangas Project, Project 2201207, October 26, 2022.
· ALS<br> Kamloops, Metallurgical Testwork on Composites from the Carangas Project, New Pacific Metals<br> Corp, Project KM6848, May 31, 2023.
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· ALS<br> Kamloops, Comminution Testwork on Composites from the Carangas Project, New Pacific Metals<br> Corp, Project KM7065, September 8, 2023.
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To support the PEA in 2024, three composite samples from the oxidized domain, transitional domain and fresh (sulfide) domain in the upper silver zone and one composite sample from the lower gold zone were identified and collected in December 2023. The samples from the upper silver zone with a targeted silver grade of 60 g/t were subjected to the flotation testing to produce a silver/lead concentrate and a zinc concentrate. From these three domain composite samples, a life-of-mine (LOM) composite sample was prepared and subjected to the flotation testing to produce a silver/lead concentrate and a zinc concentrate. This LOM composite sample consisted of 12.5% oxidized domain, 2.5% transitional domain and 85.0% fresh (sulfide) domain with a targeted silver grade of 60 g/t. Later on, another LOM composite sample was prepared and subjected to the flotation testing to produce a silver/lead concentrate and a zinc concentrate. This second LOM composite sample represented an average mill feed composition during the first eight years of commercial production, and consisted of 17.0% oxidized domain, 3.0% transitional domain and 80.0% fresh (sulfide) domain with a targeted silver grade of 74 g/t. The composite sample from the lower gold zone with a targeted gold grade of 1.0 g/t was subjected to gravity concentration, whole-ore cyanide leach, bulk flotation, selective flotation and cyanide leach of the bulk flotation concentrate. A simulation of the gravity concentration circuit was completed by FLSmidth in 2024. Three reports were issued out of this testwork program with relevant results included in the discussion below.

· FLSmidth,<br> New Pacific Metals Carangas Project Gravity Circuit Modelling Report, March 28, 2024.
· ALS<br> Kamloops, Metallurgical Testwork on Composites from the Carangas Project to Support a Pre-Feasibility<br> Study, New Pacific Metals Corp, Project KM7100, Progress Report, August 14, 2024.
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· ALS<br> Kamloops, Metallurgical Testwork on Composites from the Carangas Project to Support a Pre-Feasibility<br> Study, New Pacific Metals Corp, Project KM7100, Final Report, August 14, 2025.
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Table 13-1:         Summaryof the Completed Metallurgical Testwork

Year Laboratory Testwork
2022 Bureau Veritas<br><br>Minerals, Canada Five composite samples (silver/lead shallow zone with no oxidation, silver/lead/zinc shallow zone with partial oxidation, silver/lead/zinc deep zone with no oxidation, lower gold zone with low sulfur and lower gold zone with high sulfur). Flotation and whole-material cyanide leach. The testwork was carried out for NPM.
2023 ALS Kamloops,<br><br> Canada Three composite samples (silver/lead shallow zone with no oxidation, silver/lead/zinc shallow zone with partial oxidation and silver/lead/zinc deep zone with no oxidation). Flotation and cyanide leach of silver/lead concentrate. The testwork was conducted for NPM.
2023 ALS Kamloops,<br><br> Canada Three composite samples (upper silver zone, lower silver zone and lower gold zone). Specific gravity, abrasion index, Bond rod mill work index and Bond ball mill work index. The testwork was performed for NPM.
2024 - 2025 ALS Kamloops,<br><br> Canada Six composite samples (USZ oxidized, USZ transitional, USZ sulfide, USZ LOM1, USZ LOM2, LGZ LOM). Mineralogy, flotation, whole-material cyanide leach, cyanide leach of gold flotation concentrate, gravity concentration and cyanide leach of silver/lead concentrate. The testwork was completed for NPM.
13.2 Historical Metallurgical Testwork (2022–2023)
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13.2.1 Testwork of Flotation and Cyanide Leach (2022–2023)
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Preliminary metallurgical testwork was started in June 2022 and completed in May 2023 with five composite samples which were selected from the coarse assay sample rejects based on the mineralization and oxidation. The scope of work included the detailed head analysis, bulk flotation to generate a silver/lead concentrate, sequential selective flotation to produce a silver/lead concentrate and a zinc concentrate, whole-material cyanide leach for recovery of gold and silver, and cyanide leach of the lead/silver concentrate for recovery of silver.

From June 2022 to October 2022, the metallurgical testwork was carried out by Bureau Veritas Minerals in Richmond, British Columbia, Canada. A report entitled “Metallurgical Testing for Gold, Silver, Lead and Zinc Recovery, New Pacific Metals – Carangas Project, project# 2201207” was issued on October 26, 2022. From October 2022 onward, the remaining metallurgical testwork was continued by ALS Metallurgy in Kamloops, British Columbia, Canada. ALS Metallurgy issued a report on May 31, 2023 with a title of “Metallurgical Testwork on Composites from the Carangas Project, New Pacific Metals Corp, Bolivia, project# KM6848.”

These five composite samples were prepared using intervals from three drill holes (DCAr001, DCAr0025 and DCAr0027) in the West Dome area and another three drill holes (DCAr0019, DCAr0026 and DCAr0031) in the Central Valley area. The mineralization zones, drill holes and depths of these five composite samples are shown in Table 13-2. The contents of important elements are based on the assays from Bureau Veritas Minerals and ALS Metallurgy. Pb(Ox) and Zn(Ox) represent the contents of the oxidized lead minerals and oxidized zinc minerals, which were measured by ammonium acetate digestion.

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· Sample<br> 1 (silver/lead shallow – fully oxidized) was located nearly the surface of the deposit.<br> It contained silver and lead without zinc and gold. Sixty-nine percent of lead mineralization<br> was oxidized in this sample.
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· Sample<br> 2 (silver/lead/zinc shallow – partially oxidized) was also located nearly the surface<br> of the deposit. It contained silver, lead and zinc without gold. Thirty nine percent of lead<br> mineralization was oxidized in this sample. The oxidation of zinc mineralization was negligible.
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· Sample<br> 3 (silver/lead/zinc deep – not oxidized) was located deep in the deposit. It contained<br> silver, lead and zinc without gold. The in-situ oxidation was absent for this sample.
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· Sample<br> 4 (gold low sulfur) was located in the gold mineralization zone. It contained gold and silver<br> with sulfur content less than 1.2%.
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· Sample<br> 5 (gold high sulfur) was similar to Sample 4, but its sulfur content was higher (over 1.2%).<br> The higher sulfur content means more pyrite and might indicate the level of refractory nature<br> for gold extraction.
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Table 13-2:            Drillholes and Intervals of Five Composite Samples in 2022

Sample No. Sample Description Area Drill Hole# From To Selected Intervals
m m m
Sample 1 Silver/Lead shallow - <br><br>fully oxidized West Dome DCAr0001 23.90 43.00 14.8
DCAr0025 39.70 65.50 19.8
DCAr0027 29.00 40.50 9.7
Sample 2 Silver/Lead/Zinc shallow - <br><br>partially oxidized Central Valley DCAr0019 18.83 43.25 23.1
DCAr0026 8.00 45.87 17.1
Sample 3 Silver/Lead/Zinc deep - <br><br>not oxidized Central Valley DCAr0019 110.05 171.69 38.0
Sample 4 Gold Low-Sulfur (<1.2% S) Central Valley DCAr0031 425.28 754.48 41.2
Sample 5 Gold High-Sulfur (>1.2% S) Central Valley DCAr0031 429.69 746.30 56.8

Source: NPM, 2022

Table 13-3:            HeadAssays of Five Composite Samples in 2022

Element Unit Lab Sample 1 Sample 2 Sample 3 Sample 4 Sample 5
Ag/Pb shallow -fully oxidized Ag/Pb/Zn shallow - partially oxidized Ag/Pb/Zn deep - not oxidized Gold Low Sulfur (<1.2% S) Gold High Sulfur (>1.2% S)
Silver Ag ppm BV 199 143 157 10 8
ALS 150 100 135 - -
Lead Pb % BV 1.13 0.85 0.85 0.05 0.08
ALS 1.13 0.89 0.86 - -
Pb(Ox) % ALS 0.78 0.35 - - -
Zinc Zn % BV 0.02 0.51 1.42 0.02 0.03
ALS 0.01 0.48 1.26 - -
Zn(Ox) % ALS - 0.01 - - -
Gold Au ppm BV <0.005 <0.005 0.03 1.82 4.02
Antimony Sb ppm BV 154 52 99 28 39
Arsenic As ppm BV 194 472 184 110 154
Carbon C % BV <0.02 0.03 1.04 1.07 1.20
Copper Cu ppm BV 99 147 667 806 1,134
Mercury Hg ppm BV 0.29 0.03 0.07 0.02 0.03
Sulfur S % BV 0.41 0.70 1.84 0.62 3.07
ALS 0.39 0.70 1.88 - -

Source: Bureau Veritas Minerals, 2022; ALS Kamloops, 2023a

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Key metallurgical performances of these five composite samples are as follows:

· Cyanide<br> leach of the two gold mineralized samples (1.8–4.0 g/t Au) resulted in over 97% Au<br> recovery.
· The<br> “silver/lead shallow – fully oxidized” sample (167 g/t Ag, 1.18% Pb) contained<br> 69% of oxidized lead mineral. A high-grade silver/lead concentrate was generated from a locked<br> cycle flotation test. The concentrate contained 7,788 g/t Ag and 41.6% lead with corresponding<br> recoveries of 79.7% Ag and 60.1% Pb. One indicative cyanide leach test showed that 96% Ag<br> in this concentrate was leachable in cyanide solution in 24 hours.
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· The<br> “silver/lead/zinc shallow – partially oxidized” sample (95 g/t Ag, 0.85%<br> Pb, 0.48% Zn) contained 39% of oxidized lead mineral. A high-grade silver/lead concentrate<br> was also generated from a locked cycle flotation test, but this concentrate was significantly<br> diluted by zinc. Furthermore, it was not possible to produce a marketable zinc concentrate.<br> The silver/lead concentrate contained 5,612 g/t Ag, 35.2% Pb and 13.0% Zn with corresponding<br> recoveries of 91.1% Ag and 64.2% Pb. One indicative cyanide leach test demonstrated that<br> 94% Ag in this concentrate was leachable in cyanide solution in 24 hours.
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· The<br> “silver/lead/zinc deep – not oxidized” sample (143 g/t Ag, 0.84% Pb, 1.27%<br> Zn) corresponds to a main part of the deposit. The in-situ oxidation was absent for this<br> sample. The sequential selective flotation worked well to produce two high-quality concentrates.<br> The locked cycle flotation test generated a silver/lead concentrate containing 8,596 g/t<br> Ag and 52.1% Pb with 90.9% Ag recovery and 94.1% Pb recovery, and a zinc concentrate containing<br> 284 g/t Ag and 53.3% Zn with 3.8% Ag recovery and 80.4% Zn recovery. One indicative cyanide<br> leach test showed that 48% Ag in this silver/lead concentrate was leachable in cyanide solution<br> in 24 hours. The silver dissolution trend indicated that silver recovery would increase if<br> cyanide leach retention time was extended beyond 24 hours.
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13.2.2 Comminution Testwork (2023)
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The report of “ALS Kamloops, Comminution Testwork on Composites from the Carangas Project, NPM, Project KM7065, September 8, 2023” deals with the comminution testing of three composite samples. The drill holes and quarter core intervals, which were selected for the comminution testing, are presented in Table 13-4. Table 13-5 shows the results of SG, rod mill work index, ball mill work index and abrasion index for three composite samples.

The measured specific gravity values vary between 2.59 and 2.88. The rod mill work index values are between 10.1 kWh/t and 12.3 kWh/t. Based on these values, these samples are categorized as moderate hardness regarding the rod mill grinding. The ball mill work index values range from 10.7 to 12.8 kWh/t. These values indicate that these samples are characterized as average hardness with respect to the ball mill grinding. The abrasion index values are between 0.038 g and 0.075 g. These values indicates that these samples are classified as mildly abrasive.

Table 13-4:            DrillHoles and Core Intervals of Three Composite Samples for Comminution Testing

Mineralization Zone Drill Hole Number From To Length of Selected Intervals
m m m
Silver/Lead/Zinc Upper Silver Zone DCAr0003 46.09 56.41 10.32
DCAr0100 61.25 70.26 9.01
DCAr0141 60.94 72.50 11.56
Lower Silver Zone DCAr0045 110.05 119.00 8.96
DCAr0163 140.20 149.30 9.10
DCAr0182 168.55 178.00 9.45
Gold Lower Gold Zone DCAr0067 517.00 526.00 9.00
DCAr0104 415.95 424.99 9.04
DCAr0112 419.66 428.47 8.81

Source: NPM, 2023

Table 13-5:            SpecificGravity, Rod Mill Work Index, Ball Mill Work and Abrasion Index of Three Composite Samples

Mineralization Silver/Lead/Zinc Gold
Zone Upper Silver Zone Lower Silver Zone Lower Gold Zone
Specific Gravity 2.59 2.76 2.88
Rod Mill Work Index Feed Size (80% passing) μm 9,201 8,593 9,509
Screen Closing Size 1,180 1,180 1,180
Product Size<br><br> <br>(80% passing) 930 913 940
Rod Mill Work Index kWh/t 11.4 10.1 12.3
Ball Mill Work Index Feed Size (80% passing) μm 2,042 1,835 1,848
Screen Closing Size 106 106 106
Product Size<br><br> <br>(80% passing) 71 67 69
Ball Mill Work Index kWh/t 12.8 10.7 12.7
Abrasion Index g 0.075 0.038 0.048

Source: ALS Kamloops, 2023b

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13.3 Recent Metallurgical Testwork (2024–2025)
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A metallurgical testwork program was started in early 2024 by ALS Metallurgy in Kamloops, British Columbia, Canada, with the purpose of supporting the PEA study in 2024. For the samples from the upper silver zone, the scope of work included the detailed head assays, mineralogy, flotation to generate a silver/lead concentrate and a zinc concentrate, cyanide leach of the silver/lead concentrate to dissolve silver. For the sample from the lower gold zone, the scope of work included the detailed head assays, gravity concentration, whole-ore cyanide leach, bulk flotation, selective flotation and cyanide leach of the bulk flotation concentrate.

13.3.1 Sample Selections and Head Assays

Nine groups of samples from the upper silver zone (USZ) and one group of samples from the lower gold zone were selected (Table 13-6) from a large number of drill holes and intervals. The locations of the selected intervals for the “USZ Oxidized,” “USZ Transitional,” “USZ Fresh (Sulfide),” and “LGZ LOM” samples are shown graphically in Figure 13-1. The coarse assay sample rejects were used in each case. Five composite samples were prepared for the Upper Silver Zone (USZ), namely:

· The<br> “USZ Oxidized” composite was a silver/lead/zinc mineralization in the oxidized<br> domain with silver grade targeted at 61 g/t. Contents of lead, zinc and sulfur were<br> expected to be 0.44%, 0.08% and 0.20%, respectively.
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Figure 13-1:         Locationsof the Selected Intervals for Metallurgical Samples

Source: NPM, 2024

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Table 13-6:          TenGroups of Samples from the Upper Silver Zone and Lower Gold Zone

Mineralization Domain Sample ID Targeted HeadGrade Number of DrillHoles Number ofIntervals Total Length ofIntervals (m)
Silver, Lead and Zinc Oxidized USZ Oxidized ~61 g/t Ag 46 112 139.0
USZ Oxidized A ~120 g/t Ag 14 19 23.3
USZ Oxidized B ~30 g/t Ag 16 17 20.8
Transitional USZ Transitional ~59 g/t Ag 45 108 132.4
USZ Transitional A ~120 g/t Ag 13 16 19.9
USZ Transitional B ~30 g/t Ag 14 16 19.5
Sulfide USZ Sulfide ~60 g/t Ag 55 107 134.0
USZ Sulfide A ~120 g/t Ag 10 14 17.0
USZ Sulfide B ~30 g/t Ag 10 11 13.7
Gold - LGZ LOM ~1.01 g/t Au 23 89 115.1

Source: NPM, 2023

· The<br> “USZ Transitional” composite was a silver/lead/zinc mineralization in the transitional<br> domain with silver grade targeted at 61 g/t. Contents of lead, zinc and sulfur were expected<br> to be 0.48%, 0.65% and 0.89%, respectively.
· The<br> “USZ Fresh (Sulfide)” composite was a silver/lead/zinc mineralization in the<br> sulfide domain with silver grade targeted at 59 g/t. Contents of lead, zinc and sulfur were<br> expected to be 0.42%, 0.89% and 1.75%, respectively.
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· The<br> “USZ LOM” was a life-of-mine (LOM, 17 years) composite average mill feed with<br> 60 g/t silver, which consisted of 12.5% “USZ Oxidized”, 2.5% “USZ Transitional”<br> and 85.0% “USZ Sulfide.”
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· The<br> “USZ LOM2” was a second life-of-mine (LOM2, first 8 years) composite average<br> mill feed with 74 g/t silver, which consisted of 17.0% “USZ Oxidized”, 3.0% “USZ<br> Transitional” and 80.0% “USZ Sulfide.”
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One composite sample, LGZ LOM, was prepared for the lower gold zone, which was a LOM average mill feed with 1.01 g/t Au. Contents of silver, copper and sulfur were expected to be 11 g/t, 0.060% and 3.07%, respectively.

The contents of key elements in these six composite samples were analyzed after compositing and homogenization were completed. These assay values are shown in Table 13-7. The “PbOx” represents the amount of the oxidized lead minerals, and the “ZnOx” represents the amount of the oxidized zinc minerals, both of which were determined by ammonium acetate digestion.

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Table 13-7:         HeadAssays of the Six Composite Samples in 2024

Mineralization Domain Sample ID Au Ag Pb PbOx Zn ZnOx ST S^2-^ Cu
g/t g/t % % % % % % %
Silver, Lead and Zinc Oxidized USZ Oxidized <0.01 69 0.47 0.12 0.09 0.01 0.20 0.02 0.0084
Transitional USZ Transitional <0.01 63 0.43 0.05 0.57 0.02 0.70 0.62 0.0102
Sulfide USZ Sulfide 0.02 61 0.40 0.03 0.74 0.01 1.85 1.84 0.0220
12.5% Oxidized + 2.5% Transitional + 85.0% Sulfide USZ LOM - 58 0.41 - 0.60 - 1.50 - -
17.0% Oxidized + 3.0% Transitional + 80.0% Sulfide USZ LOM2 - 71 0.43 - 0.63 - 1.51 - -
Gold - LGZ LOM 1.03 9 0.10 0.01 0.12 <0.01 3.33 3.31 0.0566

Source: ALS Kamloops, 2025

13.3.2 Mineralogy of the Oxidized Silver/Lead/Zinc Mineralized Samples

A Particle Mineral Analysis (PMA) by QEMSCAN was completed on four size fractions of the USZ Oxidized composite sample and USZ Transitional composite sample to characterize the lead and zinc minerals at grind size of 80% passing 70 µm. Due to in-situ oxidation, these two composite samples were expected to be problematic with respect to flotation performance. The mineral composition of these two samples is shown in Table 13-8. The deportments of lead and zinc are also displayed in Figure 13-2 and Figure 13-3, respectively.

Table 13-8:          MineralCompositions of the USZ Oxidized and USZ Transitional Composite Samples

Mineral USZ Oxidized (wt%) USZ Transitional (wt%)
Copper Sulfide <0.1 <0.1
Galena <0.1 0.4
Cerussite 0.1 <0.1
Lead Sulfate 1.1 0.5
Lead Phosphate 0.1 0.2
Lead Oxide 0.3 -
Sphalerite <0.1 0.7
Zinc Oxide 0.1 0.4
Pyrite 0.1 0.7
Iron Oxide 3.3 4.6
Quartz 44.9 39.9
Feldspar 36.0 32.4
Mica 12.4 16.6
Carbonate 0.1 1.9
Titanium Mineral 0.3 0.3
Apatite <0.1 0.1
Kaolinite 0.3 0.4
Others 0.8 0.7
Total 100 100

Source: ALS Kamloops, 2024

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Figure 13-2:       Deportmentof Lead Minerals for the USZ Oxidized and USZ Transitional Composite Samples

Source: ALS Kamloops, 2024

Figure 13-3:       Deportmentof Zinc Minerals for the USZ Oxidized and USZ Transitional Composite Samples

Source: ALS Kamloops, 2024

Lead mineral was measured primarily as galena for the USZ Transitional composite sample, with a less percentage as lead sulfate and lead phosphate minerals. Approximately two thirds of the lead in the USZ Oxidized composite sample was measured as lead sulfate, lead oxide, and lead phosphate minerals. These non-sulfide lead minerals might be difficult to recover by flotation, even after sulphidizing conditioning. Cerussite, a lead carbonate mineral which will respond well to the sulphidizing conditioning for improved flotation recovery, was measured about one fifth of the lead content for the USZ Oxidized composite sample. Poor lead flotation recovery could be expected with the USZ Oxidized composite sample.

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Zinc mineral was measured primarily as sphalerite with the USZ Transitional composite sample, although one quarter of the zinc was measured within iron oxide and as zinc oxide which might not be expected to be recoverable by flotation. For the USZ Oxidized composite sample, which contained only 0.09% zinc, about 80% of the zinc was measured as zinc oxide minerals.

Quartz, feldspars and micas were the predominant silicate minerals in the USZ Oxidized and USZ Transitional composites samples. Pyrite content was 0.7% in the USZ Transitional composite sample compared to 0.1% in the USZ Oxidized composite sample. Kaolinite content was low between 0.3% and 0.4%.

13.3.3 Bulk Flotation of the USZ Oxidized Composite Sample
13.3.3.1 Rougher Flotation of the USZ Oxidized Composite Sample
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Because of the very low contents of sulfide and zinc for the USZ Oxidized composite sample, the bulk flotation procedure was applied to generate a silver/lead concentrate. Copper sulfate (CuSO4·5H2O) was used as an activator. Three collectors, namely AP3418A, A404 and SIPX, were tested. Based on prior experience, soda ash (Na2CO3) was better than lime for the pH adjustment, and thus the soda ash was used for initial three rougher flotation tests. Subsequently, the use of lime was tried again for the pH adjustment. Sulphidizing conditioning was applied to enable the oxidized lead minerals to be floatable. In total, seven rougher tests were completed. The operating conditions are shown in Table 13-9 and the results are presented in Table 13-10. Because silver is most important economically, the silver recovery was plotted against the silver content in the concentrate (Figure 13-4).

Table 13-9:         OperatingConditions of Rougher Flotation Tests for the USZ Oxidized Composite Sample

Test No. pH Lime Na2CO3 CuSO4·5H2O Redox NaHS Collector Flotation Time
AP3418A A404 SIPX
g/t g/t g/t mV g/t g/t min
03R 9.0–10.1 - 553 200 -250 2,940 20 20 - 14
07R 9.8 - added 50 -370 4,705 100 100 - 4
08R 7.8 - 294 50 -155 352 100 100 40 4
71R 8.2 added - 50 -155 352 30 30 - 4
72R 8 135 - 50 219 - - 30 30 4
77R 8.0–8.2 added - 100 -85 352 120 120 120 14
79R 8.2–9.5 added - 100 -182 to -201 1,408 120 120 - 14

Note: 1.7 kg sample, 8-liter flotation cell, grind size 80% passing ~75 µm. Source: ALS Kamloops, 2025

Figure 13-4 indicates that Tests 08R, 71R and 79R generated the best results for silver with the silver recovery between 76.9% and 83.7%, and the silver content in the concentrate between 1,082 g/t and 3,390 g/t. Lead recovery was low, between 23.4% and 41.2%. The low lead recovery is consistent with the findings from the QEMSCAN PMA investigation. The enrichment ratio between silver and lead was around 2 or above 2 except Test 72R (Figure 13-5). This indicates that silver was more favourably floated than lead. Without sulphidizing conditioning and addition of AP3418A collector (Test 72R), the lead recovery was negligible (Table 13-10).

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Table 13-10:       Resultsof Rougher Flotation Tests for the USZ Oxidized Sample

Test No. Stream Mass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
03R Concentrate 4.0 858 3.79 0.28 0.95 50.0 31.9 12.9 14.8
Feed - 69 0.48 0.09 0.26 -
07R Concentrate 3.8 476 2.81 0.20 0.65 32.8 23.4 7.2 9.6
Feed - 54 0.45 0.10 0.26 -
08R Concentrate 2.0 2,244 6.20 0.36 1.11 76.9 26.3 6.4 10.3
Feed - 59 0.48 0.12 0.22 -
71R Concentrate 1.6 3,390 10.40 0.72 1.35 79.0 35.5 14.0 10.1
Feed - 67 0.46 0.08 0.21 -
72R Concentrate 0.6 5,970 1.08 0.70 1.55 62.4 1.5 5.2 4.8
Feed - 59 0.43 0.08 0.20 -
77R Concentrate 7.0 675 2.34 0.22 0.72 80.0 41.3 28.8 24.1
Feed - 59 0.40 0.05 0.21 -
79R Concentrate 5.1 1,082 3.53 0.24 0.74 83.7 41.2 15.7 15.3
Feed - 66 0.44 0.08 0.25 -

Source: ALS Kamloops, 2025

Figure 13-4:       Relationshipbetween Silver Recovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ Oxidized Composite Sample

Source: JJ Metallurgical Services Inc, 2026

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Figure 13-5:       Silver/LeadEnrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ Oxidized Composite Sample

Source: JJ Metallurgical Services Inc, 2026

13.3.3.2 Cleaner Flotation of the USZ Oxidized Composite Sample

One open-circuit cleaner flotation test (Test 81Cl) was completed where the rougher stage followed the conditions of Test 79R. The resultant rougher concentrate was upgraded without regrinding in two stages at pH 9.0–9.1 with collector addition of 5 g/t AP3418A and 2 g/t A404 in each stage. The final concentrate at 0.57% mass pull contained 8,330 g/t silver and 27.5% lead with corresponding recoveries of 74.8% for silver and 33.7% for lead from a feed containing 63.1 g/t silver and 0.46% lead.

The concentrate after one-stage upgrade contained 4,351 g/t silver and 14.6% lead at 1.1% mass pull with corresponding recoveries of 77.7% for silver and 35.7% for lead. The rougher concentrate contained 1,208 g/t silver and 4.3% lead at 4.3% mass pull with 82.2% silver recovery and 39.9% lead recovery.

13.3.4 Sequential Selective Flotation of the USZ Transitional Composite Sample
13.3.4.1 Rougher Flotation of the USZ Transitional Composite Sample
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Six sequential selective rougher flotation tests were conducted for the USZ Transitional composite sample. The AP3418A collector was applied to float the silver/lead. The pH was adjusted using soda ash or hydrated lime. Sulphidizing conditioning was applied in two tests to a targeted redox potential. Because the zinc was poorly rejected while the silver/lead were floated with the use of zinc sulfate as a depressant, sodium metabisulfite (SMBS) was tried in Test 76R and 78R as a substitute for zinc sulfate to reject zinc. The zinc circuit followed a traditional procedure where copper sulfate was added as an activator at high pH, and the zinc was then floated with the SIPX collector. The operating conditions of these six rougher flotation tests are shown in Table 13-11, and the results are presented in Table 13-12.

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Table 13-11:       OperatingConditions of Rougher Flotation Tests for the USZ Transitional Sample

Test No. Grinding Silver/Lead Rougher
ZnSO4.7H2O SMBS Na2CO3 Ca (OH)2 pH SMBS Na2CO3 Ca (OH)2 Redox NaHS AP3418A Flotation Time
g/t g/t g/t g/t g/t g/t g/t mV g/t g/t min
04R 1,000 - - 300 9.6–9.8 - - added -253 2,940 5 14
06R 1,000 - 150 - 9.4–9.6 - added - -250 1,911 15 14
70R 1,000 - - - 9.0–9.2 - - 534 150–190 - 18 14
75R 2,000 - - - 9.0 - - 500 122–164 - 18 14
76R - 500 - - 9.0 - - 352 116–142 - 18 14
78R - - - - 6.1–6.6 1,659 382 - 161–234 - 18 14
Test No. Zinc Rougher
--- --- --- --- --- ---
pH Ca (OH)2 CuSO4·5H2O SIPX Flotation Time
g/t g/t g/t min
04R 11.0 1,729 300 38 8
06R 11.0 1,106 600 90 8
70R 11.0 1,365 600 150 8
75R 11.0 1,245 600 150 8
76R 11.0 1,246 600 150 8
78R 11.0 1,600 600 150 8

Note: 1.7 kg sample, 8-liter flotation cell, grind size 80% passing ~69 µm. Source: ALS Kamloops, 2025

Table 13-12:        Resultsof Rougher Flotation Tests for the USZ Transitional Composite Sample

Test No. Stream Solid Mass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
04R Silver/Lead Concentrate 3.0 868 7.47 2.35 14.2 49.3 51.5 12.1 55.3
Zinc Concentrate 2.2 840 2.83 3.87 2.99 34.2 14.0 14.2 8.3
Feed - 53 0.44 0.59 0.78 - - - -
06R Silver/Lead Concentrate 6.1 610 4.51 2.05 7.63 63.6 61.6 19.3 59.0
Zinc Concentrate 3.6 414 1.48 4.68 2.66 25.6 12.0 26.2 12.2
Feed - 58 0.44 0.65 0.79 - - - -
70R Silver/Lead Concentrate 5.0 1,049 5.92 6.50 9.96 88.1 63.2 52.3 66.7
Zinc Concentrate 7.9 20 0.20 2.01 1.51 2.7 3.4 25.4 15.9
Feed - 60 0.47 0.62 0.75 - - - -
75R Silver/Lead Concentrate 7.2 712 3.87 4.04 7.27 86.4 63.6 51.6 67.4
Zinc Concentrate 6.4 24 0.20 2.40 1.94 2.6 2.9 27.0 15.9
Feed - 60 0.44 0.57 0.78 - - - -
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Test No. Stream Solid Mass Composition Recovery
--- --- --- --- --- --- --- --- --- --- ---
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
76R Silver/Lead Concentrate 8.1 784 3.56 5.37 7.34 91.1 68.8 75.7 78.9
Zinc Concentrate 7.9 12 0.17 0.50 0.64 1.4 3.3 6.8 6.7
Feed - 70 0.42 0.58 0.76 - - - -
78R Silver/Lead Concentrate 5.0 931 4.90 7.82 11.74 86.4 66.3 72.1 78.2
Zinc Concentrate 13.3 12 0.14 0.53 0.44 2.8 5.2 12.9 7.7
Feed - 54 0.37 0.55 0.75 - - - -

Source: ALS Kamloops, 2025

Sulphidizing conditioning (-250 mV) for Test 04R and 06R was detrimental to silver recovery. Without sulphidizing conditioning (Test 70R, 75R, 76R and 78R), silver recovery was between 86.4% and 91.1%, and lead recovery was between 63.2% and 68.8%. However, a significant amount of zinc was floated into the silver/lead concentrate. When zinc sulfate was used as a depressant for zinc in the silver/lead circuit, about 52% zinc was floated into the silver/lead concentrate (Test 70R, 75R). When SMBS was used as a depressant for zinc in the silver/lead circuit, 72–76% zinc was floated into the silver/lead concentrate.

Figure 13-6 shows the relationship between silver recovery and silver content in the silver/lead rougher concentrate. Among these six rougher flotation tests, Test 70R was superior with respect to silver flotation performance, which achieved 88.1% silver recovery ,with the silver content in the silver/lead concentrate of 1,049 g/t (Table 13-12). Based on the trendline in Figure 13-6, silver recovery is expected to achieve 85%, when silver content in the silver/lead concentrate increases to 2,000 g/t.

Without sulphidizing conditioning (Test 70R, 75R, 76R and 78R), silver was preferentially floated over lead (Figure 13-7). However, a significant amount of zinc was floated into the silver/lead concentrate (Figure 13-17). When zinc sulfate was replaced with SMBS as a depressant for zinc in the silver/lead circuit (Test 76R and 78R), more zinc was floated into the silver/lead concentrate.

In the zinc circuit, zinc recovery was low. For Test 70R, the zinc rougher concentrate contained 2.0% zinc and 20 g/t silver with corresponding recoveries of 25.4% for zinc and 2.7% for silver.

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Figure 13-6:        SilverRecovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ Transitional Sample

Source: JJ Metallurgical Services, 2026

Figure 13-7:       Silver/LeadEnrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ Transitional Sample

Source: JJ Metallurgical Services, 2026

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Figure 13-8:        Silver/ZincEnrichment Ratio of the Silver/Lead Concentrate for the USZ Transitional Sample

Source: JJ Metallurgical Services, 2026

13.3.4.2 Cleaner Flotation of the USZ Transitional Composite Sample

One open-circuit cleaner test (Test 82) was completed for the USZ Transitional composite sample. The operating conditions in the rougher stage were similar to rougher test 70R.

· Primary<br> grinding – 1,000 g/L ZnSO4·7H2O, 80% passing 69 µm
· Silver/lead<br> rougher flotation – pH 9.0 (lime), 18 g/t AP3418A, 4 g/t Aero 404, 14 minutes
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· Regrinding<br> was not applied. The silver/lead rougher concentrate was conditioned with 500 g/t ZnSO4·7H2O
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· Silver/lead<br> 1^st^ cleaner flotation – pH 9.0 (lime), 8 g/t AP3418A, 5 minutes, the tailing<br> was forwarded to the zinc rougher
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· Silver/lead<br> 2^nd^ cleaner flotation – pH 9.0 (lime), 5 g/t AP3418A, 4 minutes, the tailing<br> was forwarded to the zinc rougher
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The operating conditions for zinc flotation were as follows:

· Zinc<br> rougher flotation – pH 11.0 (lime), 600 g/t CuSO4·5H2O,<br> 150 g/t SIPX, 8 minutes
· Regrinding<br> – 300 g/t CuSO4·5H2O, 80% passing 24 µm
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· Zinc<br> 1^st^ cleaner flotation – pH 11.0 (lime), 100 g/t SIPX, 4 minutes
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· Zinc<br> 2^nd^ cleaner flotation – pH 11.0 (lime), 50 g/t SIPX, 3 minutes
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The final silver/lead concentrate contained 3,440 g/t silver and 18.8% lead with corresponding recoveries of 84.1% for silver and 60.7% for lead. Zinc recovery (47.6%) was achieved into the silver/lead concentrate.

The final zinc concentrate contained 262 g/t silver and 43.7% zinc grades with corresponding recoveries of 1.4% for silver and 23.0% for zinc.

Table 13-13:         Resultsof the Cleaner Flotation Test for the USZ Transitional Composite Sample

Product SolidMass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
Final Silver/Lead Concentrate 1.50 3,440 18.8 19.7 32.7 84.1 60.7 47.6 66.8
Final Zinc Concentrate 0.33 262 1.04 43.7 26.8 1.4 0.7 23.0 11.9
Zinc 2^nd^ Cleaner Tail 1.22 32 0.29 0.79 0.80 0.6 0.8 1.6 1.3
Zinc 1^st^ Cleaner Tail 7.15 17 0.20 0.28 0.30 2.0 3.1 3.2 2.9
Zinc Rougher Tail 89.8 8 0.18 0.17 0.14 11.8 34.7 24.6 17.1
Feed - 61.4 0.47 0.62 0.74 - - - -

Source: ALS Kamloops, 2025

13.3.4.3 Locked Cycle Flotation of the USZ Transitional Sample

One locked cycle test (Test 84) with six cycles was completed for the USZ Transitional composite sample with operating conditions similar to the cleaner Test 82. The average results of Cycle V and VI are presented in Table 13-14. The results of this locked cycle test are similar to the open-circuit cleaner Test 82 for both silver/lead concentrate and zinc concentrate. The final silver/lead concentrate contained 3,070 g/t silver and 19.4% lead with 83.8% silver recovery and 65.4% lead recovery. Zinc recovery(45.6%) was achieved into the silver/lead concentrate. The final zinc concentrate contained 208 g/t silver and 45.1% zinc with 1.2% silver recovery and 21.1% zinc recovery.

Table 13-14:        Resultsof the Locked Cycle Test 84 for the USZ Transitional Composite Sample

Product SolidMass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
Silver/Lead Concentrate 1.50 3,070 19.4 20.1 32.1 83.8 65.4 45.6 66.3
Zinc Concentrate 0.30 208 0.64 45.1 26.9 1.2 0.4 21.1 11.4
Final Tail Zinc 1^st^ Cleaner Tail 8.50 9.0 0.06 0.20 0.20 1.4 1.1 3.0 2.8
Zinc Rougher Tail 89.6 8.4 0.16 0.22 0.16 13.7 33.0 30.3 19.5
Flotation Feed - 55.4 0.45 0.67 0.73 - - - -

Source: ALS Kamloops, 2025

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13.3.5 Sequential Selective Flotation of the USZ Sulfide Composite Sample
--- ---
13.3.5.1 Rougher Flotation of the USZ Sulfide Composite Sample
--- ---

Three sequential selective rougher tests were conducted for the USZ Fresh (Sulfide) composite sample. The collectors, AP3418A and A404, were applied. The pH was adjusted using soda ash in one test, and hydrated lime in other two tests. Sulphidizing conditioning was not included. The addition of sodium cyanide was tried to depress zinc and pyrite in the silver/lead circuit. The zinc circuit followed a traditional procedure where copper sulfate was used as an activator at high pH. The conditions of these three rougher flotation tests are shown in Table 13-15 and the results are presented in Table 13-16.

In the silver/lead circuit, the addition of sodium cyanide in Test 05R improved the rejection of pyrite and zinc, but silver recovery suffered (Table 13-16, Figure 13-11). The high slurry pulp density in Test 02R resulted in poor concentrate quality due to entrainment of gangue minerals (Figure 13-11). Test 10R generated the best results for the silver/lead concentrate, which contained 799 g/t silver and 5.44% lead with corresponding recoveries of 92.5% for silver and 85.5% for lead. The rejection of zinc in the silver/lead circuit was relatively strong, and the rejection of pyrite was apparent. Lead was floated into the silver/lead concentrate at a nearly same ratio as silver (Figure 13-10). Lead recovery was not impacted by cyanide addition in Test 05R.

In the zinc circuit, zinc flotation performance was reasonably good. Test 02R resulted in 67.2% zinc recovery at 16.8% concentrate mass pull, which corresponds to 86% stage zinc recovery. For Test 05R, zinc recovery was 82.1% at 9.2% concentrate mass pull, which corresponds to 91% stage zinc recovery.

Table 13-15:       OperatingConditions of Rougher Flotation Tests for the USZ Sulfide Composite Sample

Test No. Grinding Silver/Lead Rougher
ZnSO4.7H2O NaCN Cell Volume pH Na2CO3 Ca (OH)2 Collector Flotation Time
AP3418A A404
g/t g/t L g/t g/t g/t min
02R 1,000 - 4 9.0 - 624 9 - 8
05R 120 40 8 9.0 117 - 9 - 8
10R 1,000 - 8 9.0 - 358 18 4 10
Test No. Zinc Rougher
--- --- --- --- --- --- ---
Cell Volume pH Ca (OH)2 CuSO4·5H2O SIPX Flotation Time
L g/t g/t g/t min
02R 4 11.0 1,100 300 60 16
05R 8 11.0 1,176 300 60 16
10R -

Note: 1.7 kg sample, grind size 80% passing ~71 µm. Source: ALS Kamloops, 2024

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Table 13-16:       Resultsof Rougher Flotation Tests for the USZ Fresh (Sulfide) Composite Sample

Test No. Steam Solid Mass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
02R Silver/Lead Concentrate 10.8 514 3.63 1.64 7.5 87.9 89.7 22.0 42.2
Zinc Concentrate 16.8 30 0.09 3.21 5.5 8.0 3.6 67.2 48.4
Tailing 72.4 4 0.04 0.12 0.3 4.1 6.7 10.8 9.5
Feed - 63 0.44 0.80 1.9 -
05R Silver/Lead Concentrate 3.0 1,402 12.16 2.56 12.9 75.6 86.7 9.9 19.6
Zinc Concentrate 9.5 121 0.21 6.60 15.4 20.9 4.9 82.1 75.4
Tailing 87.5 2 0.04 0.07 0.1 3.5 8.4 8.0 4.9
Feed - 55 0.42 0.77 2.0 -
10R Silver/Lead Concentrate 7.0 799 5.44 2.65 16.2 92.5 85.5 24.4 59.9
Tailing 93.0 5 0.07 0.62 0.8 7.5 14.5 75.6 40.1
Feed - 61 0.45 0.76 1.9 -

Source: ALS Kamloops, 2024

Figure 13-9:       SilverRecovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ Sulfide Composite Sample

Source: JJ Metallurgical Services, 2026

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Figure 13-10:     Silver/LeadEnrichment Ratio of the Silver/Lead Concentrate for the USZ Sulfide Composite Sample

Source: JJ Metallurgical Services, 2026

Figure 13-11:     Silver/ZincEnrichment Ratio of the Silver/Lead Concentrate for the USZ Sulfide Composite Sample

Source: JJ Metallurgical Services, 2026

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13.3.5.2 Cleaner Flotation of the USZ Sulfide Composite Sample
--- ---

Thirteen cleaner flotation tests were completed with the USZ Sulfide composite sample at a primary grind size of 80% passing approximately 71 µm. The test variables included the pH, regrind time, collector dosage, depressant dosage, and cleaner circuit configuration. The zinc circuit was included only in the initial five cleaner tests. Subsequent cleaner tests were carried out to improve the flotation performance in the silver/lead circuit. The operating conditions of these thirteen cleaner tests are presented in Table 13-17 and the corresponding results are shown in Table 13-18.

Figure 13-12 shows the relationship between silver recovery and silver content in the silver/lead concentrate. Among these thirteen cleaner tests, Tests 90 and 91 achieved the best results for silver, followed by Test 73.

· Test<br> 91 generated a silver/lead concentrate which contained 2,130 g/t silver and 14.7% lead with<br> corresponding recoveries of 87.3% for silver and 83.1% for lead.
· Although<br> Test 90 had a slightly lower silver recovery (83.2%) than Test 91, but the lead recovery<br> was higher (87.5%) and the concentrate grade was also improved for silver and lead (3,610<br> g/t silver and 25.2% lead).
--- ---
· Test<br> 73 was a first cleaner test, which achieved fairly good results. The concentrate contained<br> 4,880 g/t silver and 33.5% lead with 80.1% silver recovery and 86.2% lead recovery.
--- ---

Figure 13-13 shows the relationship between lead recovery and lead content in the silver/lead concentrate. As with silver flotation performance, Tests 90 and 73 achieved the best results for lead. Despite highest silver recovery achieved in Test 91, the silver/lead concentrate was diluted by an elevated amount of pyrite. Test 90 produced a silver/lead concentrate which contained 25.2% lead at 87.5% lead recovery. The results of Test 73 were similar, that is, 33.5% lead content in the silver/lead concentrate at 86.2% lead recovery. Overall, the lead was floated at a similar same rate as silver (Figure 13-14). The rejection of zinc in the silver/lead circuit was reasonably good (Figure 13-15).

Figure 13-16 shows the relationship between zinc recovery and zinc content in the zinc concentrate. Based on the average values of Tests 73, 74, 80 and 83, the zinc concentrate contained 266 g/t silver and 50.2% zinc with 4.7% silver recovery and 65.1% zinc recovery.

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Table 13-17:       OperatingConditions of the Cleaner Flotation Tests for the USZ Sulfide Composite Sample

Test No. Primary Grinding Silver/Lead Flotation
Rougher Regrinding 1^st^ Cleaner 2^nd^ Cleaner
Time ZnSO4· 7H2O pH AP3418A A3477 A404 Flotation Time Feed Time ZnSO4 ·7H2O pH AP3418A A3477 A404 CMC Flotation Time pH AP3418A A3477 A404 Flotation Time
min g/t g/t g/t g/t min min g/t g/t g/t g/t g/t min g/t g/t g/t min
Test 73 10.0 1,000 9.0 18 / 4 14 Ro.<br> Conc 5 300 9.0 8 / / / 8 9.0 2 / / 5
Test<br> 74 10.0 1,000 9.0 18 / 4 14 Ro.<br> Conc 5 300 9.0 8 / / / 8 9.0 2 / / 5
Test<br> 80 10.0 1,000 9.0 18 / 4 14 Ro.<br> Conc 5 300 9.0 8 / / / 8 9.0 2 / / 5
Test<br> 83 10.0 1,000 9.0 18 / 4 14 Ro.<br> Conc 5 100 9.0 8 / / / 8 9.0 2 / / 5
Test<br> 85 10.0 1,000 9.0 18 / 4 14 Ro.<br> Conc 5 100 9.0 8 / / / 8 9.0 2 / / 5
Test<br> 86 10.0 1,000 9.0 18 18 4 14 Ro.<br> Conc 5 100 9.0 8 8 / / 8 9.0 2 2 / 5
Test<br> 87 10.0 1,000 9.0 18 / 4 14 Ro.<br> Conc 5 300 9.0 8 / / 10 8 9.0 2 / / 5
Test<br> 88 10.0 1,000 9.0 18 / 4 14 Ro.<br> Conc 5 300 9.0 10 / 5 / 8 9.0 4 / 2 5
Test<br> 89 10.0 1,000 8.5 18 / 4 14 Ro.<br> Conc 5 300 8.5 10 / 5 / 8 8.5 4 / 2 5
Test 90 10.0 1,000 8.5 18 / 4 14 1^st^<br> Cl Conc 3.5 300 8.5 2 / 2 / 8 8.5 10 / 5 5
Test 91 10.0 1,000 8.5 18 / 4 14 1^st^<br> Cl Conc 1.5 300 8.5 2 / 2 / 8 8.5 10 / 5 5
Test<br> 92 10.0 1,000 8.5 18 / 4 14 1^st^<br> Cl Conc 1.5 300 8.5 2 / 2 / 8 8.5 10 / 5 5
Test<br> 93 10.0 1,000 8.5 18 / 4 14 1^st^<br> Cl Conc 3.5 300 8.5 2 / 2 / 8 8.5 10 / 5 5
Test No. Zinc Flotation
--- --- --- --- --- --- --- --- --- --- --- --- ---
Rougher Regrinding 1^st^ Cleaner 2^nd^ Cleaner
pH CuSO4·5H2O AP3418A Flotation Time Time CuSO4·5H2O pH AP3418A Flotation Time pH AP3418A Flotation Time
g/t g/t min min g/t g/t min g/t min
Test 73 10.0 125 16 6 5 60 11.0 12 3 11.0 2 2
Test 74 10.0 125 13 6 5 60 11.0 5 3 11.0 3 2
Test 80 10.0 125 8 6 5 60 11.0 3 3 11.0 2 2
Test 83 10.0 125 8 6 5 60 11.0 3 3 11.0 2 2
Test 85 10.0 125 8 6 5 60 11.0 3 3 11.0 2 2

Source: ALS Kamloops, 2025

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Table 13-18:       Resultsof Cleaner Flotation Tests for the USZ Sulfide Composite Sample

Test No. Product Mass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
Test 73 Silver/Lead 2^nd^ Cleaner Conc 0.95 4,880 33.5 7.8 32.7 80.1 86.2 9.5 16.3
Zinc 2^nd^ Cleaner Conc 0.85 168 0.32 53.0 34.3 2.5 0.7 57.7 15
Feed - 57.8 0.37 0.78 1.90 - - - -
Test 74 Silver/Lead 2^nd^ Cleaner Conc 0.97 4,630 32.8 8.4 29.8 77.0 84.2 10.6 15.2
Zinc 2^nd^ Cleaner Conc 1.14 262 0.93 45.6 37.4 5.1 2.8 67.6 22
Feed - 58.5 0.38 0.77 1.90 - - - -
Test 80 Silver/Lead 2^nd^ Cleaner Conc 0.89 4,990 35.3 8.2 28.3 76.5 80.1 9.5 13.7
Zinc 2^nd^ Cleaner Conc 0.97 382 1.52 50.0 37.1 6.4 3.8 63.5 20
Feed - 58.1 0.39 0.77 1.84 - - - -
Test 83 Silver/Lead 2^nd^ Cleaner Conc 1.00 4,180 30.7 8.6 30.4 77.9 67.5 11.7 16.7
Zinc 2^nd^ Cleaner Conc 1.01 252 0.91 52.1 34.5 4.7 2.0 71.5 19
Feed - 53.7 0.46 0.73 1.83 - - - -
Test 85 Silver/Lead 2^nd^ Cleaner Conc 1.22 3,612 27.6 9.1 30.9 78.0 84.7 14.4 20.4
Zinc 2^nd^ Cleaner Conc 0.90 272 1.16 34.1 31.1 4.4 2.6 39.9 15
Feed - 56.3 0.40 0.77 1.84 - - - -
Test 86 Silver/Lead 2^nd^ Cleaner Conc 1.43 3,160 21.7 9.1 33.7 82.9 78.4 16.7 27.6
Feed - 54.6 0.40 0.78 1.75 - - - -
Test 87 Silver/Lead 2^nd^ Cleaner Conc 1.04 4,226 30.3 8.2 25.1 79.8 83.4 10.9 14.4
Feed - 55.1 0.38 0.78 1.82 - - - -
Test 88 Silver/Lead 2^nd^ Cleaner Conc 1.37 3,210 23.3 7.0 26.7 80.3 84.0 13.0 19.8
Feed - 54.6 0.38 0.74 1.84 - - - -
Test 89 Silver/Lead 2^nd^ Cleaner Conc 1.45 3,010 21.2 6.0 34.5 78.3 80.7 10.7 27.4
Feed - 55.9 0.38 0.81 1.83 - - - -
Test 90 Silver/Lead 2^nd^ Cleaner Conc 1.33 3,610 25.2 7.9 33.6 83.2 87.5 14.0 24.2
Feed - 57.7 0.38 0.75 1.84 - - - -
Test 91 Silver/Lead 2^nd^ Cleaner Conc 2.16 2,130 14.7 5.3 39.4 87.3 83.1 15.1 46.6
Feed - 52.6 0.38 0.76 1.82 - - - -
Test 92 Silver/Lead 2^nd^ Cleaner Conc 1.46 2,930 21.2 6.9 38.0 82.3 81.5 13.4 30.3
Feed - 51.9 0.38 0.75 1.83 - - - -
Test 93 Silver/Lead 2^nd^ Cleaner Conc 1.35 3,208 23.2 7.4 36.3 81.7 82.8 13.1 26.6
Feed - 53.1 0.38 0.76 1.85 - - - -
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Figure 13-12:     SilverRecovery and Silver Content in the Silver/Lead Concentrate for the USZ Sulfide Sample

Source: JJ Metallurgical Services, 2026

Figure 13-13:     LeadRecovery and Lead Content in the Silver/Lead Concentrate for the USZ Sulfide Sample

Source: JJ Metallurgical Services, 2026

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Figure 13-14:     Silver/LeadEnrichment Ratio of the Silver/Lead Concentrate for the USZ Sulfide Sample

Source: JJ Metallurgical Services, 2026

Figure 13-15:     Silver/ZincEnrichment Ratio of the Silver/Lead Concentrate for the USZ Sulfide Sample

Source: JJ Metallurgical Services, 2026

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Figure 13-16:     ZincRecovery and Zinc Content in the Zinc Concentrate for the USZ Sulfide Sample

Source: JJ Metallurgical Services, 2026

13.3.6 Sequential Selective Flotation of the USZ LOM Composite Sample

A LOM composite sample was prepared according to the 2024 mine production schedule over the 17 years of mine life. The LOM composite consisted of 12.5% USZ Oxidized sample, 2.5% USZ Transitional sample and 85.0% USZ Sulfide sample with silver grade target at 60 g/t. This LOM composite was subjected to a series of rougher flotation tests, open-circuit cleaner flotation tests, closed circuit cleaner flotation tests and finally the locked cycle tests.

13.3.6.1 Rougher Flotation of the USZ LOM Composite Sample

Ten rougher flotation tests were carried out with a focus on the silver/lead concentrate to investigate a series of parameters which are listed below. Details of the operating conditions are presented in Table 13-19, and the results are shown in Table 13-20.

· Depressants<br> in the silver/lead circuit to reject zinc and pyrite – (1) zinc sulfate, (2) zinc<br> sulfate plus sodium cyanide, and (3) sodium metabisulfite (SMBS)
· A<br> sulphidizing conditioning step to the 5^th^ rougher stage (Test 13R)
--- ---
· pH<br> adjustment using soda ash or hydrated lime
--- ---
· Lower<br> pH when SMBS was used
--- ---
· An<br> alternative collector, known as X5000, which was supplied by Ecolab (former Flottec).
--- ---
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Table 13-19:       OperatingConditions of Rougher Flotation Tests in the Silver/Lead Circuit for the USZ LOM Composite Sample

Test No. Grinding Silver/Lead Rougher
ZnSO4·7H2O NaCN SMBS pH Redox NaCN SMBS Na2CO3 Ca (OH)2 Collector Flotation Time
AP3418A X5000 A404
g/t g/t g/t mV g/t g/t g/t g/t g/t min
13R - - - 8.3–9.9 145/-300 - - - - 38 - 23 12
23R - - - 8.4 - - - - - - 33 18 8
28R 1,000 - - 9.0 - - - - 447 18 - 4 8
29R 1,000 - - 9.0 - - - 764 - 18 - 4 8
30R 1,000 30 - 9.1 - - - - 411 18 - 4 8
31R - - 500 5.5 - - - - - 18 - 4 8
32R - - 500 7.5 - - - - - 18 - 4 8
33R - - - 7.3 - - 500 - - 18 - 4 8
34R 1,000 30 - 9.0 - - - - 341 40 - 10 14
38R 1,000 - - 9.0 - 10 - - 176 40 - 10 14

Note: 1.7 kg sample, 8-liter flotation cell, grind size 80% passing ~71 µm. Source: ALS Kamloops, 2024

Table 13-20:       Resultsof Rougher Flotation Tests in the Silver/Lead Circuit for the USZ LOM Composite Sample

Key Chemicals Test No. Stream Solid Mass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
AP3418A, no ZnSO4 13R Silver/Lead Concentrate 9.0 603 3.68 5.32 16.7 94.9 76.9 72.6 89.3
Feed - 57 0.43 0.66 1.7 -
X5000, no ZnSO4 23R Silver/Lead Concentrate 8.4 630 4.00 4.96 17.5 93.4 78.6 59.4 87.9
Feed - 57 0.43 0.70 1.7 -
AP3418A + ZnSO4 28R Silver/Lead Concentrate 5.0 1,157 6.67 4.44 20.8 90.1 77.9 31.6 63.6
Feed - 65 0.43 0.71 1.6 -
29R Silver/Lead Concentrate 6.2 919 5.55 3.86 20.6 92.1 78.6 34.3 72.8
Feed - 62 0.44 0.70 1.8 -
Average Silver/Lead Concentrate 5.6 1,038 6.11 4.15 20.7 91.1 78.3 32.9 68.2
AP3418A + ZnSO4 + NaCN 30R Silver/Lead Concentrate 3.9 1,286 7.85 1.99 8.9 81.2 74.5 11.5 21.4
Feed - 62 0.42 0.68 1.6 -
34R Silver/Lead Concentrate 9.5 536 3.59 1.14 5.4 84.2 77.4 16.7 29.7
Feed - 61 0.44 0.65 1.7 -
38R Silver/Lead Concentrate 8.7 603 3.97 2.32 5.4 82.8 76.0 28.0 27.0
Feed - 64 0.46 0.72 1.7 -
Average Silver/Lead Concentrate 7.4 808 5.14 1.82 6.5 82.7 76.0 18.7 26.0
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Key Chemicals Test No. Stream Solid Mass Composition Recovery
--- --- --- --- --- --- --- --- --- --- --- ---
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
AP3418A + SMBS 31R Silver/Lead Concentrate 7.9 731 4.00 7.07 20.2 91.8 77.4 80.2 91.1
Feed - 63 0.41 0.70 1.8 -
32R Silver/Lead Concentrate 6.6 852 5.13 6.14 19.7 92.1 78.5 58.5 78.7
Feed - 61 0.43 0.70 1.7 -
33R Silver/Lead Concentrate 7.8 768 4.45 4.84 17.9 91.9 79.0 53.2 78.3
Feed - 65 0.44 0.71 1.8 -
Average Silver/Lead Concentrate 7.4 783 4.53 6.02 19.3 91.9 78.3 63.9 82.7

Source: ALS Kamloops, 2024

Figure 13-17 shows the relationship between silver recovery and silver content in the silver/lead rougher concentrate. Among these ten rougher flotation tests, Tests 28R and 29R achieved the best results overall. Silver recovery was similar between soda ash and hydrated lime when used for pH adjustment (Test 28R vs Test 29R). The average silver/lead concentrate of these two rougher flotation tests contained 1,038 g/t silver and 6.11% lead with 91.1% silver recovery and 78.3% lead recovery (Table 13-20). Zinc (32.9% recovery) was floated into the silver/lead concentrate.

Test 13R was carried out without zinc sulfate as a depressant. Although silver recovery and lead recovery were good (94.9% for silver and 76.9% for lead), excessive zinc (72.6% recovery) was floated into the silver/lead concentrate.

With the addition of sodium cyanide (Tests 30R, 34R and 38R), zinc rejection was good with only 18.7% (average recovery) was floated into the silver/lead concentrate. The use of sodium cyanide, either added to the grinding or to the conditioning, was detrimental to silver recovery. Silver recovery was much lower (82.7% on average) when cyanide was added. The lead recovery (76.0% on average) was not affected by sodium cyanide.

The use of SMBS as a depressant for zinc (Tests 31R, 32R and 33R) was not successful, although final silver recovery and lead recovery remained good. However, SMBS slowed down silver flotation kinetics. Average recoveries of these three rougher flotations were 91.9% for silver and 78.3% for lead. Zinc (63.9% recovery) was floated into the silver/lead concentrate.

The alternative collector, X5000, did not show any benefits compared with the collector AP3418A (Test 13R vs Test 23R).

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Figure 13-17:     SilverRecovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ LOM Sample

Source: JJ Metallurgical Services, 2024

Figure 13-18:         Silver/LeadEnrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ LOM Sample

Source: JJ Metallurgical Services, 2024

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Figure 13-19:     Silver/ZincEnrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ LOM Sample

Source: JJ Metallurgical Services, 2024

The silver/lead enrichment ratio of the silver/lead concentrate is shown in Figure 13-18 as a function of silver recovery. Overall, lead was floated consistently with silver at a silver/lead enrichment ratio between 1.0 and 1.3.

The silver/zinc enrichment ratio of the silver/lead concentrate is presented in Figure 13-19 as a function of silver recovery. Together with the silver recovery, Tests 28 and 29 were considered to be best where zinc sulfate was used as a depressant and AP3418A was used as a collector.

13.3.6.2 Cleaner Flotation of the USZ LOM Composite Sample

Eight cleaner tests were completed for the USZ LOM composite sample. Among these eight cleaner tests:

· Four<br> tests (Tests 37, 41, 42 and 56) covered only the silver/lead circuit
· Two<br> cleaner tests (Tests 49 and 50) included both silver/lead circuit and zinc circuits. Only<br> the silver/lead rougher tail was forwarded to the zinc circuit
--- ---
· Two<br> cleaner tests (Tests 51 and 52) contained both silver/lead circuit and zinc circuits. The<br> rougher tail, 1^st^ cleaner tail and 2^nd^ cleaner tail in the silver/lead<br> circuit were forwarded to the zinc circuit.
--- ---

Table 13-21 shows the operating conditions for these eight cleaner tests. The primary grinding was carried out for 10 minutes with 1.7 kg sample and an addition of 1,000 g/L ZnSO4·7H2O. The grind size was about 80% passing 68 µm. The silver/lead rougher flotation was carried out at pH 9.0 with the addition of 18 g/t AP3418A and 4 g/t A404 for 14 minutes. One issue encountered in the zinc circuit was the high slurry viscosity when pH was raised with the addition of hydrated lime. It was suspected that some gangue minerals reacted with hydrated lime at high pH. In order to move the testwork forward, sodium hydroxide was used instead in the zinc circuit. When the high slurry viscosity issue is resolved later on, the hydrated lime will be tested again in the zinc circuit because the hydrated lime is significantly cheaper than sodium hydroxide for the future commercial operations.

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Table 13-21:       OperatingConditions of the Cleaner Flotation Tests for the USZ LOM Composite Sample

Test No 37 41 42 56 49 50 51 52
Circuit Ag/Pb only Ag/Pb + Zn Ag/Pb+Zn closed circuit
Ag/Pb Rougher Ca(OH)2 g/t 241 276 294 - - -
Na2CO3 g/t - - 687 -
NaOH g/t - 300 335 - 300
Ag/Pb Regrinding Regrind Time min 5 No 3 10 10 10
ZnSO4·7H2O g/t 300 - 300 300 300 300
Ag/Pb<br><br> <br>1^st^ Cleaner ZnSO4·7H2O g/t - 300 - - - -
pH 9.0 9.5 9.0 9.0
Ca(OH)2 g/t 71 88 71 - - -
Na2CO3 g/t - - 176 - -
NaOH g/t - 185 64 - 180 185
AP3418A g/t 5 5 5
A404 g/t - 2.5 - -
Flotation Time min 5 7 5 7 7
Ag/Pb<br><br> <br>2^nd^ Cleaner pH 9.0 9.2 9.0 9.0
AP3418A g/t 2 2 2
A404 g/t - 1 - -
Flotation Time min 4 4 4
Ag/Pb<br><br> <br>3^rd^ Cleaner pH 9.0 - 9.1 9.0 -
AP3418A g/t 1 - 1
Flotation Time min 3 - 3
Zn Rougher pH - 11.0 10.2 11.0 11.5
Na2CO3 g/t - 3,811 -
NaOH g/t 1,011 - 1,100 1,752
CuSO4·5H2O g/t 100 100
SIPX g/t 40 40 5
Flotation Time min 6 6
Zn Regrinding Regrind Time min - 8 8 8 7
CuSO4·5H2O g/t 50 50
Zn 1^st^ Cleaner pH - 11.0 10.8 11.0 11.5
Na2CO3 g/t - 1,763 -
NaOH g/t 164 - 202 294
SIPX g/t 15 15 1
Flotation Time min 2 3
Zn 2^nd^ Cleaner pH 11.0 10.8 11.0 11.5
SIPX g/t 10 10 1
Flotation Time min 2 2

Note: primary grinding 1,000 g/L ZnSO4·7H2O and 10 minutes. Ag/Pb rougher at pH 9.0, 18 g/t AP3418A, 4 g/t A404 and 14 minutes.

Source: ALS Kamloops, 2025

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Table 13-22:         Resultsof the Cleaner Flotation Tests for the USZ LOM Composite Sample

Test No. Product Solid Mass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
37 Ag/Pb Conc 3^rd^ Cleaner Conc 0.64 7,050 39.3 9.2 26.8 71.7 63.2 8.8 10.5
2^nd^ Cleaner Conc 0.97 4,931 27.5 9.6 26.3 76.2 67.0 14.0 15.7
1^st^ Cleaner Conc 3.91 1,302 7.2 4.0 12.8 81.5 70.7 23.4 30.9
Rougher Conc 12.3 474 2.5 2.3 10.2 93.1 77.9 43.1 77.3
Feed - 63 0.40 0.66 1.6 -
41 Ag/Pb Conc 3^rd^ Cleaner Conc 1.88 2,460 16.6 5.9 39.1 80.1 72.3 15.9 44.8
2^nd^ Cleaner Conc 2.24 2,129 14.2 5.9 38.2 82.3 73.4 18.8 52.0
1^st^ Cleaner Conc 3.14 1,548 10.2 5.0 29.6 84.2 74.4 22.5 56.7
Rougher Conc 6.0 836 5.5 3.6 18.5 87.2 76.1 31.3 68.0
Feed - 58 0.43 0.70 1.6 -
42 Ag/Pb Conc 3^rd^ Cleaner Conc 1.15 4,490 27.8 9.2 33.6 77.5 67.9 13.7 22.3
2^nd^ Cleaner Conc 1.48 3,617 22.3 8.6 31.4 80.4 70.0 16.4 26.8
1^st^ Cleaner Conc 2.52 2,192 13.4 6.3 24.9 82.7 71.5 20.4 36.1
Rougher Conc 9.0 689 4.0 2.9 15.4 92.6 76.9 33.1 79.6
Feed - 67 0.47 0.78 1.7 -
56 Ag/Pb Conc 2^nd^ Cleaner Conc 1.22 3,528 23.1 11.5 26.4 79.0 69.7 22.4 20.5
1st Cleaner Conc 3.40 1,317 8.62 4.80 13.6 82.3 72.6 26.1 29.5
Rougher Conc 10.13 489 3.10 2.26 10.7 90.9 77.8 36.7 69.0
Feed - 54 0.40 0.62 1.6 -
49 Ag/Pb Conc 3^rd^ Cleaner Conc 0.62 7,330 45.8 10.4 24.9 75.5 68.6 10.7 10.1
2^nd^ Cleaner Conc 0.79 5,942 37.0 10.3 26.1 78.1 70.6 13.5 13.6
1^st^ Cleaner Conc 2.00 2,484 15.2 7.8 20.4 82.3 73.4 25.6 26.7
Rougher Conc 8.45 665 3.9 3.1 13.4 93.0 79.8 43.2 73.9
Zn Conc 2^nd^ Cleaner Conc 0.42 162 0.25 53.2 34.2 1.1 0.2 36.4 9.3
1^st^ Cleaner Conc 0.71 136 0.24 36.1 27.5 1.6 0.4 42.4 12.8
Rougher Conc 1.99 78 0.17 13.3 16.9 2.6 0.8 43.6 22.0
Feed - 60 0.42 0.61 1.5 -
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Test No. Product Solid Mass Composition Recovery
--- --- --- --- --- --- --- --- --- --- --- ---
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
50 Ag/Pb Conc 3^rd^ Cleaner Conc 0.43 8,990 56.1 7.5 21.0 67.1 58.6 4.7 5.6
2^nd^ Cleaner Conc 0.63 6,548 40.6 9.6 22.4 72.6 62.9 8.9 8.9
1^st^ Cleaner Conc 1.51 2,939 17.9 6.9 17.2 77.7 66.1 15.1 16.3
Rougher Conc 8.03 650 3.7 2.7 13.6 91.7 72.5 31.8 69.1
Zn Conc 2^nd^ Cleaner Conc 0.80 126 0.25 43.5 31.4 1.8 0.5 51.0 15.9
1^st^ Cleaner Conc 1.82 78 0.20 20.0 17.2 2.5 0.9 53.1 19.7
Rougher Conc 4.58 47 0.16 8.5 9.2 3.8 1.8 56.7 26.5
Feed - 57 0.41 0.69 1.6 -
51 Ag/Pb Conc 2^nd^ Cleaner Conc 0.64 7,100 41.6 12.0 24.1 75.6 66.4 11.9 10.1
Zn Conc 2^nd^ Cleaner Conc 1.32 418 1.5 34.4 36.5 9.1 4.8 69.8 31.2
1^st^ Cleaner Conc 2.30 293 1.0 20.3 26.3 11.1 5.9 71.9 39.2
Rougher Conc 5.82 153 0.6 8.4 15.8 14.7 8.1 75.2 59.6
Feed - 61 0.40 0.65 1.5 -
52 Ag/Pb Conc 2^nd^ Cleaner Conc 0.59 6,810 44.6 11.9 25.9 73.9 65.7 10.4 10.2
Zn Conc 2^nd^ Cleaner Conc 0.62 562 2.9 58.7 33.2 6.5 4.6 54.4 13.8
1^st^ Cleaner Conc 0.78 540 2.8 50.1 31.4 7.8 5.4 58.4 16.4
Rougher Conc 3.29 206 0.9 13.8 20.3 12.6 7.7 68.1 44.7
Feed - 54 0.40 0.67 1.5 -

Source: ALS Kamloops, 2025

The results of these eight cleaner tests are summarized in Table 13-22. Silver recovery into the silver/lead concentrate was relatively consistent based on the relationship between silver recovery and concentrate mass pull (Figure 13-20) or based on silver recovery and silver content in the silver/lead concentrate (Figure 13-21). Based on the trendline in Figure 13-20, the silver/lead concentrate mass pull will be around 1.3% to achieve 80% silver recovery. At 80% silver recovery, silver content in the silver/lead concentrate will be around 3,500 g/t (Figure 13-21).

The lead was floated in the silver/lead concentrate at a consistent rate with silver (Figure 13-22) with the enrichment ratio between silver and lead being around 1.15. This means when silver recovery is 80%, the lead recovery would be 80%/1.15 = 69.6%. Based on the trendline in Figure 13-23, the lead content in the silver/lead concentrate at 69.6% lead recovery would be around 24-25%.

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Figure 13-20:         SilverRecovery and Silver/Lead Concentrate Mass Pull for the USZ LOM Composite Sample

Source: JJ Metallurgical Services, 2026

Figure 13-21:         SilverRecovery and Silver Content in the Silver/Lead Concentrate for the USZ LOM Composite Sample

Source: JJ Metallurgical Services, 2026

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Figure 13-22:         Silver/LeadEnrichment Ratio of the Silver/Lead Concentrate for the USZ LOM Composite Sample

Source: JJ Metallurgical Services, 2026

Figure 13-23:         LeadRecovery and Lead Content in the Silver/Lead Concentrate for the USZ LOM Composite Sample

Source: JJ Metallurgical Services, 2026.

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Figure 13-24:         Silver/ZincEnrichment Ratio of the Silver/Lead Concentrate for the USZ LOM Composite Sample

Source: JJ Metallurgical Services, 2026

At 80% silver recovery for the silver/lead concentrate, the silver/zinc enrichment ratio is around 5 based on the trendline in Figure 13-24. This implies that at 80% silver recovery, the amount of zinc floated into the silver/lead concentrate is about 80%/5 = 16%.

With respect to the zinc concentrate, its quality was good overall. The zinc content in the concentrate was 53.2% for Test 49, 43.5% for Test 50, 34.4% for Test 51 and 50.1–58.7% for Test 52. When only silver/lead rougher tail was forwarded to the zinc circuit, zinc recovery was low, 36.4% for Test 49 and 51.0% for Test 50. When silver/lead rougher tail and cleaner tails were forwarded to the zinc circuit, zinc recovery increased substantially, 69.8% for Test 51 and 54.4–58.4% for Test 52 (Table 13-22).

13.3.6.3 Locked Cycle Tests of the USZ LOM Composite Sample

Four locked cycle flotation tests were completed for the USZ LOM composite sample. The first two locked cycle tests suffered from inadequate concentrate mass pulls in the cleaner stage for silver/lead concentrate and zinc concentrate. The third locked cycle test (Test 55) was carried out to meet the concentrate mass pull targets described above, and the flotation performance improved. The fourth locked cycle test (Test 57) was carried out with further change by running the first cleaner stage in the zinc circuit in a closed loop to improve zinc recovery in the zinc concentrate.

To alleviate the issue of high slurry viscosity in the zinc circuit with the addition of lime, sodium hydroxide was used for pH adjustment in the silver/lead circuit and also in the zinc circuit. In the future, when the slurry viscosity issue is resolved, the hydrated lime will be tried again for pH adjustment. The operating conditions for the locked cycle Test 55 are presented in Table 13-23. The operating conditions for locked cycle Test 57 were same as Test 55 except that the first cleaner stage in the zinc circuit was run in a closed circuit.

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Table 13-23:            OperatingConditions of the Locked Cycle Test 55 for the USZ LOM Composite Sample

Silver/Lead Circuit Zinc Circuit
Primary<br> Grinding Zinc<br> Sulfate ZnSO4·7H2O g/t 1,000 -
Grind<br> Size (P80) µm 68 -
Rougher pH 9.0 11.5
Sodium<br> Hydroxide NaOH g/t 300 625
Collector<br> AP3418A g/t 18 -
Collector<br> A404 g/t 4 -
Copper<br> Sulfate CuSO4·5H2O g/t - 100
Collector<br> SIPX g/t - 8
Flotation<br> Time min 14 6
Regrinding Regrind<br> Time min 11 10
Regrind<br> Size (P80) µm 19 16
Zinc<br> Sulfate ZnSO4·7H2O g/t 300 -
Copper<br> Sulfate CuSO4·5H2O g/t - 50
1^st^<br> Cleaner pH 9.0 11.5
Sodium<br> Hydroxide NaOH g/t 50 75
Collector<br> AP3418A g/t 8 -
Collector<br> SIPX g/t - 3
Flotation<br> Time min 8 3
2^nd^<br> Cleaner pH 9.0 11.5
Collector<br> AP3418A g/t 2 -
Collector<br> SIPX g/t - 1
Flotation<br> Time min 5 2

Note: 1.7 kg sample each cycle. Five cycles total. Open circuit for the Ag/Pb 1^st^ cleaner. Closed circuit for the Ag/Pb 2^nd^ cleaner. 8.0-liter cell for rougher. 2.2-liter for cleaners. Source: ALS Kamloops, 2024

The results of the locked cycle Test 55 and Test 57 are shown in Table 13-24. For the silver/lead concentrate, the average results between Test 55 and Test 57 were:

· Concentrate<br> mass pull was 1.16%.
· Silver<br> recovery was 82.1%.
--- ---
· Lead<br> recovery was 74.1%.
--- ---
· The<br> silver/lead concentrate contained 3,675 g/t silver and 24.4% lead.
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Table 13-24:            Resultsof the Locked Cycle Tests 55 and 57 for the USZ LOM Composite Sample

Stream Test No. Solid Mass Concentrate Grade Recovery
Ag Pb Zn Ag Pb Zn
% g/t % % %
Head<br> Grade Test<br> 55 - 53.1 0.40 0.66 - - -
Test<br> 57 - 50.5 0.36 0.67 - - -
Silver/Lead<br> Concentrate Test<br> 55 1.16 3,671 25.1 12.7 80.2 72.5 22.3
Test<br> 57 1.15 3,678 23.6 10.3 84.0 75.7 17.6
Average 1.16 3,675 24.4 11.5 82.1 74.1 20.0
Zinc<br> Concentrate Test<br> 55 1.07 287 0.93 35.6 5.8 2.5 57.8
Test<br> 57 1.03 332 0.88 45.3 6.8 2.5 69.3

Note: The first cleaner of zinc circuit was run in an open circuit for Test 55 and in a closed circuit for Test 57. Source: ALS Kamloops, 2025

The performance of the zinc circuit was significantly improved in Test 57, when the first cleaner stage was run in a closed loop. Based on Test 57, the zinc concentrate had the following performance.

· Concentrate<br> mass pull was 1.03%.
· Silver<br> recovery was 6.8%.
--- ---
· Zinc<br> recovery was 69.3%.
--- ---
· The<br> zinc concentrate contained 332 g/t silver and 45.3% zinc.
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13.3.7 Sequential Selective Flotation of the Second USZ LOM2 Composite Sample
--- ---

A second LOM composite sample (USZ LOM2) was prepared according to the 2024 mine production schedule over the first 8 years of mine life. The USZ LOM2 composite sample consisted of 17.0% USZ Oxidized sample, 3.0% USZ Transitional sample and 80.0% USZ Sulfide sample with silver grade targeted at 74 g/t. The USZ LOM2 composite sample was subjected to a series of rougher tests, cleaner tests and a locked cycle test.

13.3.7.1 Rougher Flotation of the USZ LOM2 Composite Sample

Five rougher flotation tests were completed for the USZ LOM2 composite sample. The primary grinding was carried out for 9.5 minutes with 1.7 kg sample and an addition of 1,000 g/t ZnSO4·7H2O, which resulted in a grind size of 80% passing 75–79 µm. The best operating conditions from the USZ LOM composite samples were applied to the USZ LOM2 composite sample. Copper sulfate was used to activate zinc prior to zinc rougher flotation with SIPX as a collector. Table 13-25 shows the operating conditions of these five rougher flotation tests, and the corresponding results are summarized in Table 13-32.

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Table 13-25:            OperatingConditions of the Rougher Flotation Tests for the USZ LOM2 Composite Sample

Test No. Silver/Lead Rougher Zinc Rougher
NaOH Ca(OH)2 Collector pH NaOH Ca(OH)2 CuSO4·5H2O Collector
AP3418A A404 SIPX
g/t g/t g/t g/t g/t g/t g/t
58R 140 - 18 4 11.5 520 - 100 8.0
59R - 355 18 4 11.5 - 1,471 100 4.0
60R - 355 18 4 9.7–10.2 - 500 100 5.5
61R - ~355 4 1 10.0 - 287 50 5.5
62R - 355 9 2 10.1–10.3 - 500 50 7.5

Note: Silver/lead rougher: pH 9.0, 14 minutes. Zinc rougher: 6 minutes. Source: ALS Kamloops, 2025

Table 13-26:            Resultsof the Rougher Flotation Tests for the USZ LOM2 Composite Sample

Test No. Conc# Mass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
58R Silver/Lead<br> Concentrate 8.7 843.4 3.93 3.29 13.98 93.6 76.8 39.7 73.6
Zinc<br> Concentrate 4.2 40.8 0.17 8.72 8.54 2.2 1.6 50.6 21.6
Head - 78.2 0.44 0.72 1.65 - - - -
59R Silver/Lead<br> Concentrate 8.4 890.9 4.28 3.31 14.13 92.9 75.1 36.6 73.5
Zinc<br> Concentrate 6.7 24.0 0.14 5.44 4.12 2.0 2.0 47.7 17.0
Head - 80.8 0.48 0.76 1.62 - - - -
60R Silver/Lead<br> Concentrate 8.2 897.8 4.14 3.57 15.47 93.0 76.8 42.6 78.0
Zinc<br> Concentrate 3.2 49.8 0.14 9.95 8.81 2.0 1.0 47.0 17.6
Head - 78.9 0.44 0.68 1.62 - - - -
61R Silver/Lead<br> Concentrate 7.5 968.3 4.59 3.26 14.99 90.5 76.7 33.4 68.6
Zinc<br> Concentrate 3.0 90.9 0.19 11.58 13.10 3.4 1.3 48.1 24.3
Head - 80.0 0.45 0.73 1.63 - - - -
62R Silver/Lead<br> Concentrate 7.5 938.6 4.68 3.44 15.72 90.4 77.1 37.9 70.4
Zinc<br> Concentrate 4.4 49.3 0.17 6.39 8.61 2.8 1.6 41.5 22.8
Head - 78.1 0.46 0.68 1.68 - - - -

Source: ALS Kamloops, 2025

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Figure 13-25:          SilverRecovery and Silver Content in the Silver/Lead Rougher Concentrate for the USZ LOM2 Sample

Source: JJ Metallurgical Services Inc, 2026

Figure 13-26:          Silver/LeadEnrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ LOM2 Sample

Source: JJ Metallurgical Services Inc, 2026

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Figure 13-27:          Silver/ZincEnrichment Ratio of the Silver/Lead Rougher Concentrate for the USZ LOM2 Sample

Source: JJ Metallurgical Services Inc, 2026

Figure 13-28:          StageZinc Recovery and Zinc Content in the Zinc Rougher Concentrate for the USZ LOM2 Sample

Source: JJ Metallurgical Services Inc, 2026

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With respect to the silver/lead concentrate, the silver flotation performance was consistent for Tests 58R, 59R and 60R (Figure 13-25), all of which had an addition of 18 g/t AP3418A and 4 g/t A404. Average silver/lead concentrate of these three rougher tests contained 877 g/t silver, 4.12% lead and 3.39% zinc with corresponding recoveries of 93.2% for silver, 76.2% for lead and 39.7% for zinc, at a mass pull of 8.4%. The lead was floated consistently with silver at a silver/lead enrichment ratio around 1.2 (Figure 13-26). The rejection of zinc in the silver/lead circuit was relatively poor (Figure 13-27), which was possibly caused by a higher percentage of the oxidized material in this composite sample.

With respect to the zinc concentrate, the zinc flotation performance was noticeably variable among these five rougher flotation tests (Figure 13-28). In comparison, Test 58R and Test 60R generated better results.

· For<br> Test 58R, the zinc concentrate contained 8.72% zinc and 41 g/t silver with corresponding<br> recoveries of 50.6% for zinc and 2.2% for silver. The stage zinc recovery was 50.6%/(1 –<br> 39.7%) = 83.9%
· For<br> Test 60R, the zinc concentrate contained 9.95% zinc and 50 g/t silver with corresponding<br> recoveries of 47.0% for zinc and 2.0% for silver. The stage zinc recovery was 47.0%/(1 –<br> 42.6%) = 81.9%
--- ---
13.3.7.2 Cleaner Flotation of the USZ LOM2 Composite Sample
--- ---

Five cleaner tests were completed with the USZ LOM 2 composite sample. The collectors AP3418A and A404 were used to float the silver/lead. After activation of sphalerite with copper sulfate, the collector AP3418A collector was used to float the zinc. The primary grinding was carried out with 1.7 kg sample with an addition of 1,000 g/t ZnSO4·7H2O for 9.5 minutes, achieving a grind size of 80% passing 73–78 µm. The silver/lead rougher flotation was conducted at pH 9.0 (lime) with additions of 18 g/t AP3418A and 4 g/t A404 for a total of 14 minutes. The silver/lead rougher concentrate was then reground with an addition of 300 g/t ZnSO4·7H2O for 11 minutes for Test 63, 64 and 65; 6 minutes for Test 68 and 3 minutes for Test 69. The silver/lead 1^st^ cleaner was carried out at pH 9.0 (lime) with an addition of 8 g/t AP3418A for 8 minutes. The silver/lead 2^nd^ cleaner was performed at pH 9.0 (lime) with an addition of 2 g/t AP3418A for 5 minutes. Lime was also used in the zinc circuit for pH adjustment. Due to the issue of high slurry viscosity, the addition of lime to the zinc rougher stage was capped at 500 g/t Ca(OH)2. The zinc 2^nd^ cleaner was carried out at pH 11.0 with an addition of 2 g/t AP3418A for 2 minutes. Other variable conditions in the zinc circuit are presented in Table 13-27.

Table 13-27:            OperatingConditions of the Cleaner Flotation Tests in the Zinc Circuit for the USZ LOM2 Sample

Test No. Zinc Rougher Zinc Regrinding Zinc 1^st^ Cleaner
pH Ca(OH)2 CuSO4·5H2O AP3418A Flotation Time Time CuSO4·5H2O pH AP3418A Flotation Time
g/t g/t g/t min min g/t g/t min
63 10.1 500 100 14 6 13 50 11.0 10 3
64 10.2 500 100 14 6 8 50 11.0 10 3
65 10.3 500 125 16 6 5 50 11.0 12 3

Source: ASL Kamloops, 2025

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Table 13-28:            Resultsof the Cleaner Flotation Tests for the USZ LOM2 Sample

Test No. Product Solid<br><br> <br>Mass Composition Recovery
Ag Pb Zn S Ag Pb Zn S
% g/t % % % %
63 Silver/Lead<br> 2^nd^ Cleaner Concentrate 0.81 7,060 39.7 13.3 25.7 75.3 67.0 15.2 13.7
Silver/Lead<br> 1^st^ Cleaner Concentrate 1.95 3,062 17.0 7.3 18.3 78.7 69.0 20.2 23.5
Zinc<br> 2^nd^ Cleaner Concentrate 0.58 380 1.0 54.6 35.4 2.9 1.2 44.6 13.5
Zinc<br> 1^st^ Cleaner Concentrate 0.74 339 0.9 45.6 32.7 3.3 1.4 47.5 15.9
Feed - 76 0.48 0.71 1.52 - - - -
64 Silver/Lead<br> 2^nd^ Cleaner Concentrate 0.85 6,980 36.4 13.8 27.2 80.0 71.2 16.6 15.2
Silver/Lead<br> 1^st^ Cleaner Concentrate 1.84 3,338 17.1 8.2 22.3 83.5 73.2 21.5 27.1
Zinc<br> 2^nd^ Cleaner Concentrate 0.70 284 0.6 53.7 34.1 2.7 1.0 54.2 15.9
Zinc<br> 1^st^ Cleaner Concentrate 1.30 210 0.5 31.2 25.4 3.7 1.4 58.1 21.9
Feed - 73 0.43 0.70 1.50 - - - -
65 Silver/Lead<br> 2^nd^ Cleaner Concentrate 1.04 5,480 28.7 13.5 25. 81.7 68.1 20.2 17.7
Zinc<br> 2^nd^ Cleaner Concentrate 1.10 152 0.7 37.2 34.9 2.4 1.6 58.6 25.2
Zinc<br> 1^st^ Cleaner Concentrate 2.00 133 0.5 21.6 24.9 3.8 2.4 62.1 32.9
Feed - 70 0.44 0.69 1.51 - - - -
68 Silver/Lead<br> 2^nd^ Cleaner Concentrate 1.20 4,650 25.0 12.1 29.3 80.8 73.5 20.8 22.8
Silver/Lead<br> 1^st^ Cleaner Concentrate 3.52 1,671 8.9 5.5 16.0 84.8 76.1 27.7 36.4
Feed - 69 0.41 0.70 1.54 - - - -
69 Silver/Lead<br> 2^nd^ Cleaner Concentrate 1.10 4,960 27.1 12.1 31.5 79.7 71.0 19.1 22.4
Silver/Lead<br> 1^st^ Cleaner Concentrate 2.64 2,159 11.7 6.9 19.8 83.0 73.2 26.0 33.6
Feed - 69 0.42 0.70 1.55 - - - -

Source: ASL Kamloops, 2025

The results of these five cleaner tests are presented in Table 13-28. The silver/lead concentrate was relatively consistent except a noticeable deviation from Test 63. The 2,000 g/t silver content in the silver/lead concentrate was easily achievable at silver recovery around 84% (Figure 13-29). At 4,500 g/t silver content in the silver/lead concentrate, silver recovery would drop slightly to 81%.

· The<br> lead was floated consistently with silver with a silver/lead enrichment ratio being around<br> 1.15 (Figure 13-30). Example: when the silver recovery is 81.0%, the lead recovery will be<br> about 81.0%/1.15 = 70.4%. Assuming head grades are 71 g/t silver and 0.44% lead, and targeting<br> 4,500 g/t silver content in the silver/lead concentrate, the corresponding lead content in<br> the silver/lead concentrate would be 4,500/71/1.15 x 0.44% = 24.2% (Figure 13-31).
· The<br> rejection of zinc in the silver/lead circuit was relatively poor. At 80% silver recovery,<br> the silver/zinc enrichment ratio was about 4.2 (Figure 13-32). This implies that 80%/4.2 = 19%<br> zinc was floated into the silver/lead concentrate. The poor zinc rejection was possibly caused<br> by the higher percentage (17.0%) of the oxidized material in this composite sample.
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Figure 13-29:          SilverRecovery and Silver Content in the Silver/Lead Concentrate for the USZ LOM2 Sample

Source: JJ Metallurgical Services Inc, 2026

Figure 13-30:          Silver/LeadEnrichment Ratio of the Silver/Lead Concentrate for the USZ LOM2 Composite Sample

Source: JJ Metallurgical Services Inc, 2026

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Figure 13-31:          LeadRecovery and Lead Content in the Silver/Lead Concentrate for the USZ LOM2 Composite Sample

Source: JJ Metallurgical Services Inc, 2026

Figure 13-32:          Silver/ZincEnrichment Ratio of the Silver/Lead Concentrate for the USZ LOM2 Composite Sample

Source: JJ Metallurgical Services Inc, 2026

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Figure 13-33:          ZincRecovery and Zinc Content in the Zinc Concentrate for the USZ LOM2 Composite Sample

Source: JJ Metallurgical Services Inc, 2026

The zinc concentrate was consistent between Tests 64 and Test 65 but was somewhat poorer for Test 63 (Figure 13-33). The zinc content in the zinc concentrate after two stages of upgrade was between 37.2% and 54.6%.

· The<br> silver/lead rougher tail and 1^st^ cleaner tail were forwarded to the zinc circuit<br> for Test 64. The net zinc recovery was 54.2%, which corresponds a stage zinc recovery of<br> 54.2%/(1 – 21.5%) = 69.0%.
· For<br> Test 65, the silver/lead rougher tail, 1^st^ cleaner tail and 2^nd^ cleaner<br> tail were all forwarded to the zinc circuit. The net zinc recovery was 58.6%, which corresponds<br> a stage zinc recovery of 58.6% / (1 – 20.2%) = 73.4%.
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13.3.7.3 Locked Cycle Flotation Tests of the Second USZ LOM2 Composite Sample
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Two locked cycle flotation tests were completed for the USZ LOM2 composite sample. The conditions of cleaner Test 65 were used for the first locked-cycle test (Test 66). Zinc sulfate was used as a zinc depressant while the silver/lead were floated with the collectors of AP3418A and A404. The pH of silver/lead rougher and cleaner was adjusted to pH 9.0 using lime. After the silver/lead were floated, copper sulfate was used to activate the zinc. The pH in the zinc rougher was adjusted by lime, but the lime dosage was capped at 500 g/t Ca(OH)2 due to the overly high slurry viscosity when the lime dosage was over 500 g/t Ca(OH)2. The lime was also used in the zinc cleaner for pH adjustment. The AP3418A collector was used in the zinc circuit.

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The second locked-cycle test (Test 67) was completed to determine the effect of a higher silver/lead concentrate mass pull and coarser silver/lead regrind size on the silver/lead flotation performance. However, the silver/lead flotation performance of Test 67 was poorer than Test 66. The lower silver/lead rougher concentrate mass pull was a possible reason, and also likely the oxidation was responsible for the poorer flotation performance of Test 67. Despite a higher addition of copper sulfate in the zinc circuit for Test 67, the zinc flotation performance of Test 67 was poorer than Test 66. The oxidation of zinc sulfide mineral was probably responsible for the poorer performance in the zinc circuit. Table 13-29 shows the operating conditions of these two locked cycle tests. The corresponding results are presented in Table 13-30.

Table 13-29:            Conditionsof the Locked Cycle Tests for the USZ LOM2 Composite Sample

Circuit Silver/Lead Circuit Zinc Circuit
Test No. 66 67 66 67
Primary<br> Grinding Zinc<br> Sulfate ZnSO4·7H2O g/t 1,000 1,000 -
Grind<br> Time min 9.5 9.5
Grind<br> Size (P80) µm 76 73
Rougher pH 9.0 9.0 10.2 10.2
Lime<br> Ca (OH)2 g/t 355 355 500 500
Collector<br> AP3418A g/t 18 18 16 16
Collector<br> A404 g/t 4 4 -
Copper<br> Sulfate CuSO4·5H2O g/t - 100 125
Flotation<br> Time min 14 14 6 6
Regrinding Regrind<br> Time min 11 9 5 5
Regrind<br> Size (P80) µm 12 13 14 16
Zinc<br> Sulfate ZnSO4·7H2O g/t 300 300 -
Copper<br> Sulfate CuSO4·5H2O g/t - 50 60
1^st^<br> Cleaner pH 9.1 9.0 10.9 10.9
Lime<br> Ca (OH)2 g/t 76 76 129 129
Collector<br> AP3418A g/t 8 8 12 12
Flotation<br> Time min 8 8 3 3
2^nd^<br> Cleaner pH 9.0 9.0 11.5 11.0
Collector<br> AP3418A g/t 2 2 2 2
Flotation<br> Time min 5 5 2 2

Source: ALS Kamloops, 2025

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Table 13-30:            Resultsof the Locked Cycle Tests for the USZ LOM2 Composite Sample

Stream Test No. Solid<br><br> <br>Mass Composition Recovery
Ag Pb Zn Ag Pb Zn
% g/t % % %
Head<br> Grade Test<br> 66 - 72.7 0.45 0.73 - - -
Test<br> 67 - 67.0 0.48 0.73 - - -
Silver/Lead<br> Concentrate Test<br> 66 0.97 6,088 31.2 15.6 81.2 67.7 20.7
Test<br> 67 1.18 4,528 26.2 11.0 79.7 64.3 17.8
Zinc<br> Concentrate Test<br> 66 0.96 343 0.77 46.3 4.5 1.7 61.2
Test<br> 67 1.16 388 1.30 33.3 6.7 3.1 52.8

Source: ALS Kamloops, 2025

The silver/lead concentrate from these two locked cycle tests contained 4,528–6,088 g/t silver and 26.2–31.2% lead with corresponding recoveries of 79.7–81.2% for silver and 64.3–67.7% for lead. The zinc concentrate contained 343–388 g/t silver and 33.3–46.3% zinc with corresponding recoveries of 4.5–6.7% for silver and 52.8–61.2% for zinc.

The recovery at a given concentrate grade from the locked cycle test is usually higher than the open-circuit cleaner test. However, such an expectation was not consistently realized in these two locked cycle tests, especially for the lead recovery in the silver/lead concentrate and the zinc recovery in the zinc concentrate. Because each locked cycle test lasted for days, the issue of oxidation was likely responsible for the lower recovery in the locked cycle tests compared with the open-circuit cleaner test.

Figure 13-34 shows the comparison of silver recoveries in the silver/lead concentrate between the cleaner tests and the locked cycle tests. The silver recoveries of the locked cycle tests 66 and 67 reasonably fall on the trendline of the cleaner tests 64, 65, 68 and 69. Between these two locked cycle tests, the silver recovery of Test 66 was close to what was expected, but the silver recovery of Test 67 was poorer than was expected.

With respect to the lead recovery in the silver/lead concentrate, both locked cycle tests, especially Test 67, resulted in poorer performance than what was expected based on the trendline of cleaner tests 64, 65, 68 and 69 (Figure 13-35). The oxidation of lead minerals was likely responsible for the lower recovery.

With respect to the flotation of zinc concentrate from these two locked cycle tests, the zinc recovery of Test 66 was reasonable, but the zinc recovery of Test 67 was unexpectedly poorer than what was expected (Figure 13-36).

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Figure 13-34:    SilverRecovery in the Silver/Lead Concentrate between the Cleaner Tests and the Locked Cycle Tests for the USZ LOM2 Composite Sample

Source: JJ Metallurgical Services Inc, 202

Figure 13-35:    LeadRecovery in the Silver/Lead Concentrate between the Cleaner Tests and the Locked-Cycle Tests for the USZ LOM2 Composite Sample

Source: JJ Metallurgical Services Inc, 2026

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Figure 13-36:    ZincRecovery in the Zinc Concentrate between the Cleaner Tests and the Locked Cycle Tests for the USZ LOM2 Composite Sample

Source: JJ Metallurgical Services Inc, 2026

13.3.8 Metallurgical Testing for the Sample from the Lower Gold Zone

A composite sample (1.05 g/t gold, 10 g/t silver, 0.059% copper and 3.33% sulfur), labeled as LGZ LOM, from the lower gold zone was subjected to a series of metallurgical tests, including gravity concentration, whole-material cyanide leach, bulk flotation, selective flotation to produce a copper-bearing gold concentrate and cyanide leach of the bulk flotation concentrate.

13.3.8.1 Gravity Concentration Testing and Simulation

A three-stage gravity concentration test was carried out using a laboratory centrifugal concentrator by following the Knelson’s E-GRG procedure. The obtained results are presented in Table 13-31. Total gravity recoverable gold recovery was 67.3% at 1.47% concentrate mass pull and 48.4 g/t gold in the concentrate. The size-by-size gold recoveries are shown in Figure 13-37. Most of gold particles in the gravity concentrate were between 53 µm and 300 µm, which are considered to be moderate to coarse.

A simulation was carried out by FLSmidth (former Knelson Concentrators) for a gravity concentration circuit which treats a portion of cyclone underflow for a process plant at 500 t/h mill throughput. The simulated results are present in Table 13-32. At a primary grind size of 80% passing 100 µm, the gold recovery from a gravity concentration circuit is expected to be 41.5-43.7% when 500-700 t/h cyclone underflow is fed to a gravity concentration circuit which consists of two QS48 centrifugal concentrators. When the primary grind size is reduced to 80% passing 75 µm, the expected gold recovery from the gravity concentration circuit increases slightly to 44.1-46.4%.

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Table 13-31:            Resultsof the E-GRG Gravity Concentration Test for the LGZ LOM Composite Sample

Stage Particle Size (80% Passing) Mass Pull Gold Grade Gold Recovery
Feed Concentrate
µm µm % g/t %
1^st^ 911 1,321 0.60 34.0 19.3
2^nd^ 266 313 0.38 77.8 27.7
3^rd^ 76 122 0.49 43.3 20.3
Total<br> Concentrate 1.47 48.4 67.3
Head<br> Grade - 1.06 -

Source: ALS Kamloops, 2024

Figure 13-37:            Size-by-SizeGravity Recoverable Gold for the LGZ LOM Composite Sample

Source: JJ Metallurgical Services, 2024

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Table 13-32:            ExpectedGravity Gold Recovery from a Commercial Operation

Particle Size of Cyclone Overflow (P80) Circulating Load Treated Solid Throughput to Gravity Concentration Circuit Gold Recovery Equipment
µm % t/h %
100 33 500 41.5 2<br> x QS48 (250 t/h per unit)
47 700 43.7 2<br> x QS48 (350 t/h per unit)
75 33 500 44.1 2<br> x QS48 (250 t/h per unit)
47 700 46.4 2<br> x QS48 (350 t/h per unit)

Note: 500 t/h mill throughput. 300% recirculation load in the grinding circuit. Source: FLSmidth, 2024

13.3.8.2 Bulk Flotation Tests to Produce a Gold Concentrate

The bulk flotation approach was applied to generate a gold concentrate with a composite sample from the lower gold zone (LGZ), which contained 1.05 g/t gold, 10 g/t silver, 0.059% copper and 3.33% sulfur. Seven rougher flotation tests were completed to investigate the impact of grind size, pulp density and collector dosage on the gold flotation performance. The operating conditions are presented in Table 13-33, and the obtained results are shown in Table 13-34. Overall, gold flotation performance was consistently good even at a coarse grind size of 80% passing 266 µm. Average recoveries were 98.0% for gold and 94.7% for silver at 10.9% mass pull.

Table 13-33:            OperatingConditions of the Rougher Flotation Tests for the LGZ LOM Composite Sample

Test No. Sample<br><br> <br>Weight Grind Size (P80) Flotation Cell Volume pH Collector Dosage Flotation Time
SIPX AF208
kg µm L g/t min
01R 1.7 76 4.4 9.0 12 12 8
11R 3.4 76 8.0 8.9 12 12 8
14R 1.7 140 8.0 8.9 12 12 8
16R 3.4 76 8.0 8.9 12 12 8
17R 3.4 76 8.0 8.9 12 12 8
19R 1.7 171 8.0 8.8 12 12 8
22R 1.7 266 8.0 8.9 24 24 8

Source: ALS Kamloops, 2024

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Table 13-34:            Resultsof the Rougher Flotation Tests for the LGZ LOM Composite Sample

Test No. Solid Mass Content in the Concentrate Recovery
Gold Silver Sulfur Gold Silver Sulfur
% g/t g/t % %
01R 10.8 7.49 100 30.4 97.8 94.5 97.9
11R 11.1 5.90 86 29.2 97.3 93.0 97.3
14R 9.8 10.50 88 35.6 98.3 95.0 97.5
16R 11.1 7.59 78 31.3 97.9 95.1 97.5
17R 11.6 9.92 84 28.2 99.2 95.7 96.4
19R 11.2 8.67 76 29.4 98.2 95.1 97.9
22R 10.9 8.88 83 29.8 97.3 94.4 96.3
Average 10.9 8.42 85 30.6 98.0 94.7 97.2

Source: ALS Kamloops, 2024

13.3.8.3 Selective Flotation to Produce a Copper-Bearing Gold Concentrate

The LGZ LOM composite sample contained 0.060% copper. A portion of copper minerals may dissolve during cyanide leach and thus consume sodium cyanide. Some of the dissolved copper is expected to be loaded onto the activated carbon or reduced together with gold/silver in the Merrill Crowe process. If the dissolved copper is carried over to the electrowinning circuit, it will be plated together with gold/silver and thus contaminate the quality of gold/silver doré. Two exploratory selective flotation tests were conducted to generate a copper-bearing gold concentrate. The first flotation test targeted bulk flotation in the rougher stage and then the resultant bulk rougher concentrate was upgraded by following a selective flotation approach. The second test followed a selective flotation in both rougher stage and cleaner stage. The operating conditions of these two flotation tests are shown in Table 13-35 and the results are presented in Table 13-36**.**

Table 13-35:            OperatingConditions of the Selective Flotation Tests to Produce a Copper-Bearing Gold Concentrate

Test No. Conditions for Rougher Conditions for 3-Stage Cleaner
pH Collector Dosage Flotation Time Regrind Size (P80) pH NaCN Collector Dosage Flotation Time
SIPX AF208 AP3418A SIPX AF208 AP3418A
g/t min µm g/t g/t min
24Cl 8.9 12 12 - 8 19 11.5 20 4+3+2 4+3+2 - 5+4+3
25Cl 10.5 / 12 12 8 8 11.5 5 - 5+4+3 5+4+3 5+4+3

Note: 3.4 kg sample, grind size 80% passing 100 µm, 8.0-liter cell for rougher, 2.2-liter cell for cleaner. Source: ALS Kamloops, 2024

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Table 13-36:            Resultsof the Selective Flotation Tests to Produce a Copper-Bearing Gold Concentrate

Test No. Product Solid<br><br> <br>Mass Composition Recovery
Cu Au Ag S Cu Au Ag S
% % g/t g/t % %
24Cl 3^rd^<br> Cleaner Concentrate 0.16 9.10 350 2,272 14.8 23.5 55.2 36.6 0.7
2^nd^<br> Cleaner Concentrate 0.38 6.25 159 1,242 13.9 39.0 60.6 48.3 1.6
1^st^<br> Cleaner Concentrate 1.66 2.54 41 377 20.2 70.2 69.9 64.9 10.3
Rougher<br> Concentrate 11.2 0.52 9 83 28.4 97.0 99.1 95.4 97.5
Feed - 0.060 0.99 9.7 3.3 -
25Cl 3^rd^<br> Cleaner Concentrate 0.14 17.2 370 1,564 20.4 38.7 51.5 22.8 0.8
2^nd^<br> Cleaner Concentrate 0.27 11.5 212 1,126 16.5 51.3 58.3 32.4 1.3
1^st^<br> Cleaner Concentrate 0.78 5.55 84 616 15.2 70.7 66.3 50.7 3.4
Rougher<br> Concentrate 4.52 1.29 20 172 18.9 95.1 89.2 81.8 24.8
Feed - 0.061 0.99 9.5 3.5 -

Source: ALS Kamloops, 2024

Both flotation tests recovered a significant amount of copper. In comparison, Test 25Cl, which followed a selective flotation approach in rougher and cleaner stages, generated better results. The concentrate quality and recoveries of Test 25Cl were as follows:

· After<br> the rougher concentrate was reground and upgraded in three cleaner stages, the final concentrate<br> contained 17.2% copper, 370 g/t gold and 1,564 g/t silver with corresponding recoveries of<br> 38.7% for copper, 51.5% for gold and 22.8% for silver. This copper-bearing gold concentrate<br> is expected to be saleable as a copper concentrate.
· If<br> the reground rougher concentrate was upgraded in two cleaner stages, the concentrate quality<br> was slightly reduced. Containing 11.5% copper, 212 g/t gold and 1,126 g/t silver with corresponding<br> recoveries of 51.3% for copper 58.3% for gold and 32.4% for silver r. Although copper content<br> was reduced, this copper-bearing gold concentrate is saleable as a gold concentrate if it<br> cannot be sold as a copper concentrate.
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· If<br> the reground rougher concentrate was upgraded in one cleaner stage, the copper content in<br> the concentrate was further decreased to 5.55%, the concentrate contained 84 g/t gold and<br> 616 g/t silver with corresponding recoveries of 70.7% for copper, 66.3% for gold and 50.7%<br> for silver. This concentrate will be saleable as a gold concentrate.
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13.3.8.4 Whole-Ore Cyanide Leach of the LGZ LOM Composite Sample
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Two whole-ore cyanide leach tests were completed for the LGZ LOM composite sample. The operating conditions and results are presented in Table 13-37. Primary grind size ranged from 80% passing 90-107 µm, and cyanide concentration was between 0.75 and 1.00 g/L NaCN. The total retention time was 48 hours. The first 30-hour cyanide leach was carried out in the absence of activated carbon. Activated carbon was then added, and the cyanide leach was continued for another 18 hours. The final gold recovery varied between 91.2% and 94.0%, and silver recovery varied between 42.3% and 50.8%. Sodium cyanide (NaCN) consumption varied between 0.42 and 0.55 kg/t.

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Table 13-37:            Conditionsand Results of the Cyanide Leach of the LGZ LOM Composite Sample

Test No Grind Size (80) pH Cyanide Concentration Head Grade Recovery Reagent Consumption
Au Ag Au Ag NaCN CaO
µm g/L NaCN g/t % kg/t kg/t
21CN 107 11.0 0.75 1.13 8.7 91.2 42.3 0.42 0.93
27CN 90 12.0 1.00 0.83 10 94.0 50.8 0.55 4.35

Note: Bottle roll, 40% solid, 30-hour DCN + 18-h CIP, continuous oxygen sparging. Source: ALS Kamloops, 2024

13.3.8.5 Cyanide Leach of the Bulk Gold Flotation Concentrate

Four cyanide leach tests were completed to investigate the impact of particle size, retention time, lead nitrate and activated carbon. The operating conditions and results are shown in Table 13-38. The final gold recovery varied between 94.0% and 96.2%, and silver recovery varied between 55% and 68%. Although the back-calculated gold head grade was variable, the tail grade was relatively consistent between 0.25 g/t and 0.28 g/t for the reground concentrate. Without regrinding, the tail contained 0.48 g/t gold. These results imply that modest regrinding is helpful to gold recovery.

Table 13-38:            Conditionsand Results of the Cyanide Leach of the Bulk Gold Flotation Concentrate

Test No. Particle Size (P80) pH Retention Time Pb (NO3)2 Head Grade Recovery Reagent Consumption
Au Ag Au Ag NaCN CaO
µm h kg/t g/t % kg/t kg/t
12CN ~76 10.5-12.0 72-h<br> DCN - 9.01 70 94.7 55 3.0 3.1
20CN 8 11.0-11.6 55-h<br> DCN+17-h CIP 0.5 6.57 63 96.2 68 14.2 1.2
26CN 10 11.5-12.0 30-h<br> DCN + 18-hCIP - 4.70 70 94.0 68 8.7 1.7
36CN 15 11.2-11.5 72-h<br> DCN - 6.62 68 96.2 64 5.1 1.3

Note: Bottle roll, 25% solid, 6.0 g/L NaCN, continuous oxygen sparging. DCN – cyanide leach without activated carbon. CIP – cyanide leach with activated carbon. Source: ALS Kamloops, 2024

13.4 Metallurgical Variability

The mineralisation of Carangas deposit is divided into a silver zone on the near surface and a gold zone at depth . The silver zone is further divided into the oxidized domain, transitional domain and fresh (sulfide) domain. Since 2022, the mineralized samples from the silver zone and gold zone have been subjected to a series of metallurgical testing. The first group of five composite samples were tested for flotation and cyanide leach in 2022 and 2023. Because of their high head grades and the limited number of drill holes, the testwork results of these samples are not used in the current economic evaluation.

· The<br> 1^st^ composite sample was fully oxidized near the surface in the silver zone, which<br> is located in the West Dome area. This sample contained silver (167 g/t) and lead (1.18%)<br> without zinc. A total of 44.3 m of intervals from three drill holes (DCAr0001, DCAr0025,<br> DCAr0027) were chosen to make up this composite sample.
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· The<br> 2^nd^ composite sample was partially oxidized near the surface in the silver zone,<br> which is located in the Central Valley area. This sample contained silver (95 g/t), lead<br> (0.85%) and zinc (0.48%). 40.2 meters of intervals from two drill holes (DCAr0019, DCAr0026)<br> were selected to make up this composite sample.
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· The<br> 3^rd^ composite sample was not oxidized at a depth in the silver zone, which is<br> located in the Central Valley area. This sample contained silver (143 g/t), lead (0.84% Pb)<br> and zinc (1.27%). 38.0 meters of intervals from one drill hole (DCAr0019) were used to make<br> up this composite sample.
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· The<br> 4^th^ and 5^th^ composite samples were at a depth in the gold zone, which<br> is located in the Central Valley area. 41.2 meters of intervals from one drill hole (DCAr0031)<br> were selected for the 4^th^ sample (1.82 g/t gold, 10.0 g/t silver, 0.62% sulfur);<br> 56.8 meters of intervals from the same drill hole (DCAr0031) were chosen for the 5^th^<br> sample (4.02 g/t gold, 7.7 g/t silver and 3.07% sulfur).
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The second group of three composite samples were used for the comminution testing in 2023, including SG, rod mill work index, ball mill work index and abrasion index.

· The<br> 1^st^ composite sample consisted of 30.9 meters of intervals from three drill holes<br> (DCAr0003, DCAr0100, DCAr0141) at a depth between 46.09 meter and 72.50 meters in the silver<br> zone.
· The<br> 2^nd^ composite sample was prepared using 27.6 meters of intervals from three drill<br> holes (DCAr0045, DCAr0163, DCAr0182) at a depth between 110.05 and 178.00 meter in the silver<br> zone.
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· The<br> 3^rd^ composite sample was composited using 26.8 meters of intervals from three<br> drill holes (DCAr0067, DCAr0104, DCAr0112) at a depth between 415.95 and 526.00 meters in<br> the gold zone.
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The third group of six composite samples were tested in 2024 and 2025 for flotation, gravity concentration and cyanide leach to support the PEA study in 2024. Numerous intervals from a large number of drill holes were selected to prepare three domain composite samples in the silver zone and one domain composite sample in the gold zone. A large area was covered as shown in Figure 13-1 for the spatial distribution of the selected drill holes and intervals.

· The<br> Oxidized domain composite sample in the silver zone had the targets of 60 g/t silver, 0.41%<br> lead, far less than 0.74% zinc and 0.3–0.4% sulfur. In total, 112 intervals from 46<br> drills holes with a combined length of 130.0 meters were chosen to make up this composite<br> sample.
· The<br> Transitional domain composite sample in the silver zone was targeted at 60 g/t silver, 0.41%<br> lead, less than 0.74% zinc and about 0.70% sulfur. In total, 108 intervals from 45 drill<br> holes with a combined length of 132.4 meters were selected to prepare this composite sample.
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· The<br> Fresh (sulfide) domain composite sample in the silver zone was expected to contain 60 g/t<br> silver, 0.41% lead, 0.74% zinc and 2.00% sulfur. Altogether, 107 intervals from 55 drill<br> holes with a combined length of 134.0 meters were used to make up this composite sample.
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· The<br> targets of 1.01 g/t gold, 13 g/t silver and 2.93% sulfur were used for the domain composite<br> sample in the gold zone. Overall, 89 intervals from 23 drill holes with a combined length<br> of 115.1 meters were selected to prepare this composite sample.
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· Apart<br> from the three domain composite samples in the silver zone, an overall LOM composite sample<br> was also prepared for testing with silver grade targeted at 60 g/t, this consisted of 12.5%<br> Oxidized domain, 2.5% Transitional domain and 85.0% Fresh (sulfide) domain.
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· A<br> second life-of-mine (LOM2) composite sample was additionally prepared in the silver zone<br> with silver grade targeted at 74 g/t, which consisted of 17.0% Oxidized domain, 3.0% Transitional<br> domain and 80.0% Fresh (sulfide) domain.
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The 2026 mine production schedule indicates a LOM average blending ratio of 12.8% Oxidized domain, 3.0% Transitional domain and 84.2% Fresh (sulfide) domain. This LOM blending ratio is nearly identical to the situation in 2024. However, the LOM average grades have decreased since 2024.

· For<br> the Oxidized domain, the 2026 LOM average grades are 39.2 g/t silver, 0.44% lead and 0.28%,<br> compared with 59.6 g/t silver, 0.45% lead and 0.29% zinc in 2024.
· For<br> the Transitional domain, the 2026 LOM average grades are 41.4 g/t silver, 0.46% lead and<br> 0.75% zinc, compared with 63.4 g/t silver, 0.66% lead and 1.00% zinc in 2024.
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· For<br> the Fresh (sulfide) domain, the 2026 LOM average grades are 35.6 g/t silver, 0.32% lead and<br> 0.69% zinc, compared with 63.7 g/t silver, 0.42% lead and 0.90% zinc in 2024.
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The 2026 mine production schedule shows the LOM average grades for the gold zone are 0.74 g/t gold, 11.9 g/t silver and 2.75% sulfur. This gold grade is lower than the LGZ composite sample (~1.00 g/t gold) tested in 2024.

13.5 Deleterious Elements

The silver/lead concentrate and zinc concentrate, which were produced from the cleaner tests and locked cycle tests in 2024 and 2025, have not yet been assayed for the deleterious elements, such as arsenic, antimony, mercury, cadmium and fluoride, etc. The quotation from Trafigura on June 20, 2024 indicated the penalty charges for the deleterious elements. Four penalty charges were applied to the silver/lead concentrate, namely:

· Arsenic<br> penalty charge at US$2.0/tonne per 0.1% when arsenic content was above 0.50%.
· Antimony<br> penalty charge at US$2.0/tonne per 0.1% when antimony content was above 0.50%
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· Tin<br> penalty charge at US$2.0/tonne per 0.1% when tin content was above 0.50%
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· Bismuth<br> penalty charge at US$2.0/tonne per 0.01% when bismuth content was above 0.05%
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The 2024 Trafigura quotation also specified four penalty charges for the zinc concentrate, including:

· Arsenic<br> penalty charge at US$2.0/tonne per 0.1% when arsenic content was above 0.50%.
· Antimony<br> penalty charge at US$2.0/tonne per 0.1% when antimony content was above 0.50%.
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· Iron<br> penalty charge at US$2.0/tonne per 1.0% when iron content was above 8.0%.
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· Silica<br> penalty charge at US$2.0/tonne per 1.0% when silica content was above 3.0%.
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Table 13-39 shows the contents of arsenic, bismuth, mercury, antimony and tin in the Oxidized domain sample, Transitional domain sample and Fresh (sulfide) domain sample, which were calculated from the assay data of the selected intervals. If the silver/lead concentrate mass pull is 1.15%, the 0.50% threshold for antimony/tin, the 0.05% threshold for bismuth and the 40 ppm threshold for mercury will not be exceeded. The iron recovery from two locked cycle tests of the LOM composite sample was 6.2–7.4%. If arsenic is present as arsenopyrite, arsenic recovery is expected to be similar to the iron recovery. Under such assumption, the expected arsenic content in the silver/lead concentrate at 0.02% arsenic in the mill feed would be approximately 6.8% x 0.02%/1.15% = 0.12%. This is below the 0.50% penalty threshold.

Table 13-39:           Contentsof Arsenic, Bismuth, Mercury, Antimony and Tin in the Domain Samples Calculated from the Assay Data of the Selected            Intervals

Sample Arsenic Bismuth Mercury Antimony Tin
As Bi Hg Sb Sn
% % ppm % %
Oxidized<br> Domain Sample 0.027 0.00001 0.082 0.0022 0.00005
Transitional<br> Domain Sample 0.015 0.00001 0.075 0.0019 0.00006
Fresh<br> (Sulfide) Domain Sample 0.014 0.00012 0.065 0.0044 0.00003

Source: JJ Metallurgical Services Inc, 2026

If the zinc concentrate mass pull is 0.99%, the 0.50% threshold for antimony and the 25 ppm threshold for mercury will not be exceeded. The iron recovery from two locked cycle tests of the LOM composite samples was 3.7–5.1% for the zinc concentrate. If arsenic is present as arsenopyrite, arsenic recovery is expected to be similar to the iron recovery. Under such assumption, the expected arsenic content in the zinc concentrate at 0.02% arsenic in the mill feed would be approximately 4.4% x 0.02%/0.99% = 0.09%. This is below the 0.50% penalty threshold.

· The iron content in the zinc concentrate from two locked cycle tests of the LOM composite<br> samples was 10.3–13.3%. If the 8.0% threshold is applicable, a small penalty charge<br> (US$4.6–10.6/t) will incur.
· The<br> silica content in the zinc concentrate has not been determined yet. It is possible that the<br> silica content in the zinc concentrate may be above 3.0%.
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13.6 Recovery Estimates
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For the mineralized materials in the silver zone, metal recoveries and flotation concentrate grades are summarized in Table 13-40. The flotation testwork data with the low-grade materials are currently not available and will be generated in the future so that the relationships between flotation performance and feed head grades can be derived.

For the mineralized materials in the gold zone, average flotation recoveries achieved were 98.0% for gold and 94.7% for silver at 10.9% mass pull. Average recoveries from cyanide leach of the flotation concentrate were 95.3% for gold and 63.8% for silver. Further testwork data on flotation and cyanide leach will be generated in the future to finalize a preferred flowsheet and to maximize the value of the mineralized materials.

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Table 13-40:            MetalRecoveries and Concentrate Grades in the Silver Zone

Sample Oxidized Sample Transitional Sample Fresh (Sulfide) Sample LOM Composite (12.5% Oxide + 2.5% Transitional + 85.0% Fresh) LOM2 Composite<br><br> <br>(17.0% Oxide + 3.0% Transitional + 80.0% Fresh)
Head<br> Grade Silver g/t 63.1 55.4 57.7 51.8 72.7
Lead % 0.46 0.45 0.38 0.38 0.45
Zinc % 0.09 0.67 0.75 0.67 0.73
Silver/Lead<br> Concentrate Mass<br> Pull % 0.57 1.50 1.33 1.15 0.97
Content Silver g/t 8,330 3,070 3,610 3,675 6,088
Lead % 27.5 19.4 25.2 24.4 31.2
Zinc % 1.8 20.1 7.9 11.5 15.6
Recovery Silver % 74.8 83.8 83.2 82.1 81.2
Lead 33.7 65.4 87.5 74.1 67.7
Test<br> ID CLT81 LCT84 CLT90 LCT55,<br> 57 LCT66
Zinc<br> Concentrate Mass<br> Pull % - 0.30 0.99 1.03 0.96
Content Silver g/t - 208 266 332 343
Lead % - 0.64 0.92 0.88 0.77
Zinc % - 45.1 50.2 45.3 46.3
Recovery Silver % - 1.2 4.7 6.8 4.5
Zinc - 21.1 65.1 69.3 61.2
Test<br> ID - LCT84 CLT73,74,80,83 LCT57 LCT66

Source: ALS Kamloops, 2025

13.7 Comments on Mineral Processing and Metallurgical Testing

The mineralized materials in the silver zone and gold zone of Carangas deposit are moderate in hardness with Bond ball mill work index between 10.7 and 12.8 kWh/t and mildly abrasive with abrasion index between 0.038 and 0.075 g.

The mineralized materials in the silver zone are suitable to sequential selective flotation to produce a silver/lead concentrate and a zinc concentrate. The silver flotation performance in the silver/lead concentrate is relatively consistent, regardless of the extent of oxidation (Figure 13-38). At 60 g/t silver head grade, over 80% silver recovery is expected when the silver content in the silver/lead concentrate is 2,000 g/t.

The lead flotation performance (i.e. Recovery) into the silver/lead concentrate is sensitive to the extent of oxidation (Figure 13-39). The lead from the Oxidized material is floated poorly. The lead flotation performance is generally improved with the Transitional material. With the fresh (sulfide) material, the lead flotation performance is reasonably good, and at 0.40% lead head grade, about 85% lead recovery is expected when the lead content in the silver/lead concentrate is 24.0%.

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Figure 13-38:            SilverRecovery and Silver Content in the Silver/Lead Concentrate for the Silver Zone

Source: JJ Metallurgical Services Inc, 2026

The rejection of zinc in the silver/lead circuit is relatively poor and becomes worse when the material is oxidized. The LOM composite sample is floated well with the expected performance for silver and lead, but it becomes difficult to reject zinc in the silver/lead circuit. Zinc sulfate is used as depressant for zinc in the silver/lead circuit. Sodium cyanide is effective to depress zinc in the silver/lead circuit, but it reduces silver recovery significantly. As a result, the use of cyanide has been abandoned. Sodium metabisulfite (SMBS) has been tried as a depressant for zinc in the silver/lead circuit, but it was found not to be effective.

The zinc concentrate is floated under a standard procedure. However, due to high slurry viscosity at high pH, the addition of lime to the rougher stage has been limited to 500 g/t Ca (OH)2. The collectors SIPX and AP3418A have been tried separately in the zinc circuit, and both of them work well. When the zinc head grade is around 0.70%, the zinc recovery was slightly below 70%, with the zinc content in the concentrate is 45%.

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Figure 13-39:          LeadRecovery and Lead Content in the Silver/Lead Concentrate for the Silver Zone

Source: JJ Metallurgical Services Inc, 2026

The mineralized material in the gold zone is suitable to gravity concentration, bulk flotation and cyanide leach. The gold is largely free milling. Gold particles are relatively coarse. Over 40% gold is expected to be recovered in a commercial gravity concentration circuit. The bulk flotation achieved high recoveries for gold and silver. The cyanide leach testwork on both the bulk flotation concentrate and the mill feed resulted in satisfactory old recovery, but silver recovery is relatively poor.

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14 Mineral Resource Estimates
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14.1 Introduction
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The QP has independently completed a Mineral Resource estimate in accordance with the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and mineral reserves (CIM (2014) definitions). The Mineral Resource has an effective date of March 31, 2026, using the data provided by NPM as of June 1, 2023.

Mineral Resource estimates are not precise calculations, and are dependent on the interpretation of limited information on the location, shape, and continuity of the occurrence and the available sampling results.

Information contained in this Technical Report is based on information provided by NPM and verified, where possible, by the QP. SLR carried out all statistical analyses and Mineral Resource estimates under the supervision of the QP.

The Company has developed 3D mineralized models for Ag, Au, Zn, and Cu zones, and SLR has developed independent models and validated the Company's models through volume/geometry comparisons. SLR constructed a 3D digital estimate workflow for the Ag, Au, Zn, and Cu grades and compiled the Mineral Resource model based on the statistical analysis of the data provided. The SLR QP considers that the Mineral Resource estimate meets the general guidelines for CIM (2014) definitions for reporting of Mineral Resources at the Indicated and Inferred confidence levels.

The QP is not aware of any other factors, including environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that could materially affect the Mineral Resource estimate.

14.2 Resource Database

The primary source documents for the Mineral Resource estimate were:

· Drill<br> hole files (collar, down-hole survey, lithology, assay, RQD, core recovery, alteration, structure,<br> and mineralization) in CSV format
· SG<br> (density) measurements from drill core samples in CSV format
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· 3D<br> models for the main mineralized zones
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· An<br> orthophoto file in TIF format
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· A<br> one metre detailed topography file in SHP format.
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14.2.1 Sample Data
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A comprehensive dataset of drill hole collar, survey, assay, and geological records in digital format was provided to SLR on June 1, 2023.

The Carangas drill hole database contains 189 drill holes representing 81,145 m. A total of 58,212 samples was analyzed and comprise the current database for Mineral Resource estimation. Assays below the detection limit were assigned to one-half of the detection limit by NPM personnel.

A total of 27,170 RQD and 27,173 core recovery measurements from 189 drill holes existed in the database. The average core recovery within the modelled mineralized zone is 98%, ranging from 0% to 100%. Poor sample recovery is concentrated in the overburden zone or in cavities (historical artisanal mining or natural cavities).

The QP is of the opinion that the core recovery is acceptable for geological interpretation, modelling, and Mineral Resource classification.

14.2.2 Bulk Density Data

A total of 5,366 specific gravity measurements from 189 diamond drill holes exist from the Carangas deposit. Measurements were calculated using the weight in air versus the weight in water method (Archimedes), by applying the following formula:

The average bulk density for each block in the 3D mineralized domain was estimated within each domain separately, using hard boundaries and the inverse distance squared (ID^2^) function and considering a minimum of two samples and a maximum of four samples to estimate a block value. The estimated density values were used for tonnage calculations in the Mineral Resource Estimate. The density sample statistics for each zone are presented in Table 14-1.

Table 14-1:            DensityStatistics Table

Zone N Samples Density (t/m³) SD Minimum Maximum
Upper<br> Silver Zone 1,666 2.08 0.23 1.33 3.49
Middle<br> Zinc Zone 713 2.30 0.19 1.37 3.01
Lower<br> Gold Zone 877 2.27 0.21 1.20 3.22

Source: compiled by SLR, 2023

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14.3 Depletion Areas
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Historical artisanal mining activities are evident in the project area, although this activity was limited to a few meters below the surface. NPM did not supply 3D artisanal mining models for depletions, thus areas that have been subject to artisanal mining are included in the Mineral Resource. The QP does not envisage that this will materially influence the Mineral Resource estimate, as the artisanal mining was not extensive and only in the West Dome and East Dome areas, and not into the Central Valley, where the bulk of mineralization is located. Currently, there are no artisanal mining activities at the Carangas Project.

14.4 Geological Models

Geological interpretations of the lithological units and the geological structure were used to guide and interpret the shape of the mineralized wireframes, along with assay results.

The 3D mineralized zones were interpreted based on a silver equivalent (AgEq) variable and geological knowledge from the geology team. The AgEq formula is as follows:

The price assumptions for the metals are Ag: $41.00/oz, Au: $3,300/oz, Pb: $1.00/lb, Zn: $1.30/lb, Cu: $4.00/lb. Prices are based on bank and industry forecasts as of February 2026.

A modelling cut-off grade of 20 g/t AgEq was used to create wireframes of mineralization. Although mineralization modelling was based on this cut-off grade approach, some unmineralized material was included in the envelopes to maintain mineralization continuity. This is considered suitable for the mineralization style.

The mineralization zones were relatively continuous; however, they may terminate against or be displaced by structural features. In some areas, primarily in the down-dip direction, internal unmineralized material was included to maintain continuity.

The mineralized zones were built using Leapfrog Geo^TM^ software, considering all major lithologies and the transitions from each domain. The main modelled domains were developed for Ag, Au, Zn (PbZn), and Cu (Figure 14-1) as described below:

· Upper<br> Silver Zone (GM_Ag): generated using AgEq cut-off grade of 20 g/t. The upper boundary is<br> the bottom of the Overburn lithology, and the lower boundary is the bottom of ABC (andesitic<br> basalt).
· Middle<br> Zinc Zone (GM_PbZn): the grade shell was built using an AgEq cut-off grade of 20 g/t. The<br> upper boundary is the bottom of ABC, and the other boundaries are GM_Au and GM_Cu 3D wireframes.
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· Lower<br> Gold Zone (GM_Au): generated grade shell using Au cut-off grade of 0.14 g/t. The upper boundary<br> is the bottom of ABC.
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· Lower<br> Copper Zone (GM_Cu): This domain was generated using a Cu cut-off grade of 0.15%. The boundary<br> used was the GM_Au wireframe domain. This zone is not considered material due to a shortage<br> of drill information at the depth of the mineralization system.
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Following the geological knowledge, a variable orientation was used for the Upper Silver Zone. The variable orientation is based on andesite basalt (ABC) surface contact, which controls the ellipsoid orientation when building the 3D model on the silver domain.

The Carangas deposit is described as a suite of metallic sulfide and gangue minerals occurring in volcanic and intrusive rocks as veins/veinlets, breccia fillings, and dissemination. Mineralization is controlled by the temperature and pressure of the hydrothermal system, i.e., the depth below the ground surface or the distance from the heat source generated by rhyolitic intrusions.

Figure 14-1:            Three-DimensionalView of the Carangas Geological Model

Source: NPM, 2023

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14.5 Resource Assays
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A complete exploratory data analysis was performed. Univariate statistics, histograms, and box-whisker plots were constructed to investigate the dataset and determine grade capping and compositing requirements. A log histogram for silver composites is shown in Figure 14-2, and univariate statistics for all main grades are shown in Table 14-2.

Figure 14-2:            AgLog Histogram for 1.5 m Composites

Source: SLR, 2023

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Table 14-2:      UnivariateStatistics of Grade Composites, by Domain

Domain Variable N Sample Mean SD Minimum Maximum
Upper<br> Silver Zone Ag_ppm 15,743 47.80 203.16 0.04 14,182.00
Au_ppm 3,046 0.02 0.12 0.01 3.29
Cu_pct 15,743 0.01 0.03 0.00 1.61
Pb_pct 15,743 0.38 0.51 0.00 14.22
Zn_pct 15,743 0.68 0.86 0.00 16.76
Middle<br> Zinc Zone Ag_ppm 6,803 9.92 93.29 0.01 7,332.35
Au_ppm 4,871 0.05 0.06 0.01 0.58
Cu_pct 6,803 0.01 0.03 0.00 0.60
Pb_pct 6,803 0.29 0.33 0.00 7.39
Zn_pct 6,803 0.67 0.58 0.00 5.13
Lower<br> Gold Zone Ag_ppm 8,383 8.81 26.91 0.02 1,158.93
Au_ppm 8,349 0.82 2.15 0.01 53.98
Cu_pct 8,383 0.07 0.16 0.00 6.30
Pb_pct 8,383 0.10 0.29 0.00 10.94
Zn_pct 8,383 0.17 0.41 0.00 7.43
Lower<br> Copper Zone Ag_ppm 154 15.33 28.27 0.13 247.88
Au_ppm 140 0.07 0.06 0.01 0.34
Cu_pct 154 0.30 0.23 0.00 1.45
Pb_pct 154 0.17 0.85 0.00 9.89
Zn_pct 154 0.37 0.99 0.00 9.02

Source: compiled by SLR, 2023

14.6 Grade Capping/Outlier Restrictions

Applying high-grade cuts reduces the impact of extreme grade outliers on the grade estimate. It aims to prevent these statistical outliers from significantly affecting the Mineral Resource Estimate. The high-grade cuts applied to the composites were determined from the histograms and log probability plots for each element. A detailed domain study was completed, and the same global top-cut values were concluded. The high-grade cut values are shown in Table 14-3. Figure 14-2 shows the probability plots for Ag (g/t) and Au (g/t) with their respective outlier restrictions.

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Table 14-3:      TopCut Values into all Domains

Variable Minimum Maximum Capping Value
Ag_ppm 0.0 9,626 7,000
Au_ppm 0.0 53.977 40.0
Pb_pct 0.0 14.220 No<br> Capping
Zn_pct 0.0 16.760 No<br> Capping
Cu_pct 0.0 1.612 No<br> Capping

Source: compiled by SLR, 2023

Figure 14-3:      Agand Au Probability Plots with capping threshold

Source: SLR, 2023

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14.7 Compositing
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The information in this section addresses only the data used to estimate the Mineral Resources.

Although the most common sample length inside the mineralized wireframes was 1.28 m (Figure 14-3), a composite length of 1.5 m was selected to reduce variability and the coefficient of variation (CV). Reducing the CV during compositing reduces the risk of metal loss when applying high-grade cuts.

The composites were checked visually in Leapfrog Geo™ software for spatial correlation with the wireframed mineralized envelopes and to assess the impact of the 1.5 m composite length. The QP considered the chosen composite length to be representative of local variations.

Figure 14-4:      LengthHistogram for Raw Assay Intervals

Source: SLR, 2023

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14.8 Estimation/Interpolation Methods
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14.8.1 Block Model Strategy and Analysis
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A series of upfront test modelling was completed to define an estimation methodology to meet the following criteria:

· Representative<br> of the current Carangas geological and structural models
· Accounts<br> for the variability of grade, orientation, and continuity of mineralization
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· Controls<br> the smoothing (grade spreading) of grades and the influence of outliers
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· It<br> is robust and repeatable within the mineral domains
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· Supports<br> multiple domains.
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Multiple test scenarios were evaluated to determine the optimal processes and parameters to achieve the stated criteria. Each scenario was based on nearest neighbor (NN), ID^2^, inverse distance cubed (ID^3^), and ordinary kriging (OK) interpolation methods.

All test scenarios were evaluated based on global statistical comparisons, visual comparisons of composite assays versus block grades, and overall smoothing assessment. Based on the test results, it was determined that the final resource estimation methodology would constrain mineralization by using hard-wireframe boundaries to control the spread of high-grade and low-grade mineralization. ID^2^ was selected as the interpolation method that best represents both the current Carangas database and deposit characteristics.

14.8.2 Grade Interpolation

The ID^2^ algorithm was used to estimate grades, with hard boundaries for each domain. The NN estimation method was also used for comparison grade validation and swath plot analysis.

The estimation parameters were based on the outcomes of the geospatial analysis and reflect the interpreted variability of the underlying grade continuity. A minimum and maximum number of samples was set to limit over-smoothing. A minimum of one sample was required for estimation, and a maximum of 40 samples were used. The search quadrant sector was also applied with a maximum of 10 samples per sector. The search ellipsoid ranges were based on AgEq continuity of grades and drill grid spacing.

Search orientations for the Upper Silver Zone and Middle Zinc Zone were based on the shape of the ABC contact surface. The grade interpolation strategy and parameters are presented in Table 14-4.

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Table 14-4:        CarangasGrade Estimation Search Parameters

General Value Clipping Ellipsoid Ranges Ellipsoid Directions
Domain Numeric Values Upper Bound Maximum Intermediate Minimum Dip Dip Azimuth Pitch Variable Orientation
Upper<br> Silver Zone Ag_ppm 7,000 175 175 75 - - - Yes
Upper<br> Silver Zone Au_ppm 40 175 175 75 - - - Yes
Upper<br> Silver Zone Cu_pct - 175 175 75 - - - Yes
Upper<br> Silver Zone Pb_pct - 175 175 75 - - - Yes
Upper<br> Silver Zone Zn_pct - 175 175 75 - - - Yes
Lower<br> Gold Zone Ag_ppm 7,000 175 175 75 14.216 90.9 45.0 No
Lower<br> Gold Zone Au_ppm 40 175 175 75 14.216 90.9 45.0 No
Lower<br> Gold Zone Cu_pct - 175 175 75 14.216 90.9 45.0 No
Lower<br> Gold Zone Pb_pct - 175 175 75 14.216 90.9 45.0 No
Lower<br> Gold Zone Zn_pct - 175 175 75 14.216 90.9 45.0 No
Lower<br> Copper Zone Ag_ppm 7,000 150 150 50 10 45.0 75.0 No
Lower<br> Copper Zone Au_ppm 40 150 150 50 10 45.0 75.0 No
Lower<br> Copper Zone Cu_pct - 150 150 50 10 45.0 75.0 No
Lower<br> Copper Zone Pb_pct - 150 150 50 10 45.0 75.0 No
Lower<br> Copper Zone Zn_pct - 150 150 50 10 45.0 75.0 No
Middle<br> Zinc Zone Ag_ppm 7,000 150 150 50 - - - Yes
Middle<br> Zinc Zone Au_ppm 40 150 150 50 - - - Yes
Middle<br> Zinc Zone Cu_pct - 150 150 50 - - - Yes
Middle<br> Zinc Zone Pb_pct - 150 150 50 - - - Yes
Middle<br> Zinc Zone Zn_pct - 150 150 50 - - - Yes

Source: compiled by SLR, 2023

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14.9 Density Assignment
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As discussed in Section 14.2.2, a total of 5,366 SG measurements from 189 diamond drill holes were used to estimate the density for the resource block model, which was deemed therequired amount and distribution of SG measurements allowed for direct estimation of SG within the block model. The ID^2^ method was used and produced results comparable to those of other methods. The density interpolation strategy and parameters are presented in Table 14-5.

Table 14-5:      DensityEstimation Parameters

General Ellipsoid Ranges Ellipsoid Directions
Domain Numeric Values Maximum Intermediate Minimum Dip Dip Azimuth Pitch Variable Orientation
Upper<br> Silver Zone SG 330 195 130 - - - Yes
Middle<br> Zinc Zone SG 330 195 195 - - - Yes
Lower<br> Gold Zone SG 330 195 130 0 0 110 No
Lower<br> Copper Zone SG 380 210 210 0 0 110 No

Source: compiled by SLR, 2023

The overall results are strongly related to the density database, as expected, and the validation process shows a reasonable comparison as presented in histograms from samples and the block model in Figure 14-4. The samples histogram geometry is reproducible into the estimated blocks, and the mean and standard deviation (SD) shows a good comparison.

Figure 14-5:      EstimationDensity Histogram Validation

Source: SLR, 2023

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14.10 Block Models
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A 3D digital estimate for silver, gold, lead, zinc, and copper was constructed and the Mineral Resource model was compiled based on statistical analysis of the provided data.

A block model was created for the Carangas Project, covering the main mineralized and adjacent areas. There is no rotation for the block model, and the block sizes were selected considering the geometry of mineralization, drill grid spacing, density of assay data and selected mining unit. The block model dimensions selected were 5 m by 5 m by 5 m (X, Y, Z) with no sub-cells. The block model origins, extents and attributes are shown in Table 14-6.

Table 14-6:      CarangasBlock Model Definition Parameters

Model Parameters X Y Z
Block Model Origin 538,490 7,904,850 2810
Number of Blocks 276 210 258
Parent Block Size (m) 5 5 5
Rotation Degree No No No
Field Name Description
GM (Zone Domain) Ag –<br> Upper Silver Zone<br><br> <br>Au –<br> Lower Gold Zone<br><br> <br>PbZn –<br> Middle Zinc Zone<br><br> <br>Cu –<br> Lower Copper Zone
IJK Block<br> IJK No
XC Cell<br> Centroid – X
YC Cell<br> Centroid – Y
ZC Cell<br> Centroid – Z
XINC Cell<br> Size – X
YINC Cell<br> Size – Y
ZINC Cell<br> Size – Z
XMORIG Model<br> Origin – X
YMORIG Model<br> Origin – Y
ZMORIG Model<br> Origin – Z
NX Number<br> of Cells – X
NY Number<br> of Cells – Y
NZ Number<br> of Cells – Z
DENSITY Density
Ag_ID2 Estimated<br> Ag – Inverse distance
Ag_NS Number<br> of samples in the estimation of Ag grade
Ag_AvgD Average<br> distance of samples used in Ag estimation
Au_ID2 Estimated<br> Au – Inverse distance
Cu_ID2 Estimated<br> Cu – Inverse distance
Field Name Description
Pb_ID2 Estimated<br> Pb – Inverse distance
Zn_ID2 Estimated<br> Zn – Inverse distance
AgEq_ID2 Calculated<br> Silver equivalent
Class_Fim Resource Classification:<br><br> <br>Indicated<br><br> <br>Inferred
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14.11 Reasonable Prospects for Eventual Economic Extraction
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A RPEEE was implemented for the Carangas deposit, and the process and assumptions include:

· The<br> QP responsible for the Mineral Resource estimate is Anderson Candido, Principal Geologist<br> of SLR and Fellow AusIMM member. The QP is independent of NPM and the Carangas Project.
· The<br> effective date of the Mineral Resource estimate is March 31, 2026.
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· CIM<br> (2014) definitions were used for the Carangas Project Mineral Resource estimate.
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· Industry<br> 5-year long-term consensus average prices (consensus from long-term forecasts from banks,<br> financial institutions, and other sources) of metals as of March 2026 were used for<br> all calculations, as itemized in Table 14-7.
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· Mineral<br> Resources are not Mineral Reserves and do not have demonstrated economic viability.
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· Minor<br> variations may occur during the addition of rounded numbers.
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· The<br> pit shell was generated based on the assumptions listed in Table 14-7. These assumptions<br> were based on regional benchmarks and preliminary metallurgical data.
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· The<br> QP is not aware of any other factors regarding environmental, permitting, legal, marketing,<br> or other relevant issues which could materially affect the Mineral Resource estimate.
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Table 14-7:      CommodityPrices Used in Resource Calculation

Commodity Unit Value Assumption
Metal<br> Prices Silver<br> (Ag) $/oz 41.00
Gold<br> (Au) $/oz 3,300.00
Lead<br> (Pb) $/lb 1.00
Zinc<br> (Zn) $/lb 1.30

Source: compiled by SLR, 2026

14.11.1 Input Parameters for Resource Calculation

The cut-off grade calculation was based on the following assumptions: mining operating cost, on-site milling operating cost, tailings management facility operating cost, general and administrative (G&A) cost, royalty cost, selling cost, onsite milling metal recovery percentages, metal payable percentages, and other variables.

The cost assumptions are presented below:

· Mining<br> operating cost: US$6.00/t ($/t milled)
· On-site<br> process+tailing operating cost: US$9.00/t ($/t milled)
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· G&A<br> cost: US$3.60/t ($/t milled)
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· Royalty<br> cost: 6.0%
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· Selling<br> cost: $0.5/oz AgEq
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· Metal<br> processing recoveries percentages: 81.6% Ag, 93.4% Au, 73.4% Pb, and 66.9% Zn.
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Table 14-8:      Metalpayable percentages

Product Payable
Pb<br> Con 95.00%
Ag<br> in Pb Con 96.50%
Zn<br> Con 85.00%
Ag<br> in Zn Con 70.00%
Au<br> in Au&Ag Dore 99.80%
Ag<br> in Au&Ag Dore 95.00%

For resource cut-off calculation purposes, a mining recovery of 100% and a mining dilution of 0% were applied.

An independent maximum pit study was conducted to verify the accuracy and reproducibility of the Project’s current maximum pit. The QP is of the opinion that the current maximum pit calculation is reasonable to constrain the Mineral Resources.

14.11.2 Pit Optimization Disclaimer

The pit shell used to define the depth and extent to report the open pit Mineral Resource is preliminary. The pit shell constraint may be subject to minor change after further pit optimization study in future stages of the Project.

The pit shell-constrained Mineral Resources demonstrate reasonable prospects for eventual economic extraction (RPEEE) and highlights that the pit does not constitute a scoping study or a detailed mining study, which are required to confirm the economic viability of the Project through additional drilling and metallurgical test work. It is further noted that capital expenditures (Capital expenditure) are not included in the mining costs assumed. SLR has verified the utilized operating costs based on the Project’s databases and the processing recoveries based on the preliminary test work outlined in Section 13, along with the price noted above, in determining the appropriate cut-off grade. In conclusion, the open pit constrained Mineral Resources is considered to demonstrate RPEEE; however, additional studies and drilling are required to confirm economic viability.

14.12 Classification of Mineral Resources

The definitions of resource categories used in this report are consistent with CIM (2014) definitions and adopted by NI 43-101. In the CIM classification, a Mineral Resource is defined as “a concentration or occurrence of solid material of economic interest in or on the Earth’s crust in such form, grade or quality and quantity that there are reasonable prospects for eventual economic extraction”. Mineral Resources are classified into Measured, Indicated, and Inferred categories. A Mineral Reserve is defined as the “economically mineable part of a Measured and/or Indicated Mineral Resource” demonstrated by studies at Pre-Feasibility or Feasibility-level as appropriate. Mineral Reserves are classified into Proven and Probable categories.

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At the Carangas Project, the Mineral Resource was classified as Indicated and Inferred based on data quality, sample spacing, mineralization and grade continuity. There are currently no Mineral Reserves estimated for the Project.

As noted in the geological interpretation, mineralization varies throughout the deposit, resulting in variations in geological and grade continuity. The mineralization domains are controlled by silver, gold and zinc grade, where the silver grade is concentrated in the upper zone and gold in the lower portions. While there are grade variations observed within the closer spaced drill holes (70 m by 70 m), the deposit shows good continuity of the main mineralized zones along strike. While there is good geological continuity along strike, local variations in grade and thickness occur between current drill spacing, arising from structures and resulting in discontinuities in mineralization.

Given the likelihood of further local grade variation with further drilling, the QP considers the current data suitable to provide an estimate of tonnage and metal content on a global scale and considers the 70 m by 70 m spacing suitable for an Indicated classification. The QP is of the opinion that further drilling is required to allow for better estimates of local grade and metal distribution, and as such, no Measured Resources are reported.

The classification criteria used for the Mineral Resource was as follows:

· Indicated
o Average<br> ID^2^ sample distance of less than 70 m or NN sample distance of less than 35 m;
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o Drill<br> grid spacing of approximately 50 m to 100 m; and
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o Confirmed<br> visualization of mineralization continuity.
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· Inferred
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o Blocks<br> that do not satisfy the requirements for Indicated Resources, but had an average sample distance<br> of less than 220 m with a minimum of one drill hole, and located inside the mineralized domain<br> were classified as Inferred Resources.
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A block model view of the Mineral Resource classification is shown in Figure 14-6.

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Figure 14-6:      ClassifiedMineral Resources Block Model – Section DCAr0094

Source: SLR, 2026

14.13 Block Model Validation

The block model validation process included visual comparisons between block estimates and composite grades in section views, local versus global estimates for ID^2^ and NN, and swath plots. A three-step process was used to validate the estimation as outlined below:

· Mean<br> grade comparison in each domain
· Swath<br> plots comparing estimation methods
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· Visual<br> inspection of the blocks against drill hole composites.
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A quantitative assessment of the estimate was completed by comparing the average grades of the top-cut composite file against the block model grades for each domain. The results of the main element for each domain are tabulated in Table 14-9 and indicate a good correlation.

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Table 14-9:      Compositevs. Block Model Grade Statistical Validation

Domain Grade Sample Grade Model Grade Difference in Grade Difference (%)
Upper Silver Zone Ag (g/t) 42.39 41.20 1.19 3%
Lower Gold Zone Au (g/t) 0.77 0.75 0.02 3%
Middle Zinc Zone Pb (%) 0.29 0.30 -0.01 -3%
Middle Zinc Zone Zn (%) 0.67 0.72 -0.05 -7%
Lower Copper Zone Cu (%) 0.30 0.31 -0.01 -3%

Source: compiled by SLR, 2023

A volumetric verification was undertaken to confirm that the block model represents the mineralization wireframe volumes, and no significant discrepancy was detected. This comparison is presented in Table 14-10, indicating an excellent comparison for all mineralization veins.

Table 14-10:      3DVolumetric Model Comparison

Domain Wireframe Volume (m^3^) Model Volume (m^3^) Difference (m^3^) Difference (%)
Upper Silver Zone 80,734,000 80,760,500 -26,500 0.0%
Lower Gold Zone 52,072,000 52,063,625 8,375 0.0%
Middle Zinc Zone 37,254,000 37,242,125 11,875 0.0%
Lower Copper Zone 547,590 546,500 1,090 0.2%
Total 170,607,590 170,612,750 -5,160 0.0%

Source: compiled by SLR, 2023

Swath plots were developed to compare interpolated block grades with the sample composite data along distance slices in the X, Y, and Z directions. The swath plot analysis, shown in Figure 14-6, shows that the estimated grades had a reasonable correlation with the cut composite grades. While the swath plots preserve the overall trend between the composite and block model grades, there is variation in the composites on individual slices. This often results from the smoothing of block grades that is inherent in the Inverse Distance algorithm and the estimation parameters used. It is particularly notable when grade variations occur over short distances or when the search ellipse used for sample selection is significantly wider than the swath plot slice.

The interpolated block model's validation was assessed using visual assessments and validation plots of block grades versus capped assay grades and composites. The review demonstrated a good comparison between local block estimates and nearby assays without excessive smoothing in the block model.

Figure 14-8 and Figure 14-9 provide visual comparisons of the silver grade for the Carangas deposit. Visual comparisons for all elements were developed, and the results are acceptable. The block model grade fits the composite samples and maintains grade continuity. Overall, the visual comparison indicated that the model grades were reasonably consistent with the drill hole composite grades, both at a local scale-down dip and in areas of closer spaced drilling with grade continuity in the major direction. A reasonable degree of smoothing was observed due to a combination of the block dimensions, the ID^2^ algorithm and the wide drill spacing at some locations.

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Based on the validation results, the QP considers the estimate to be a reasonable representation of the composites and to match the known controls of mineralization and the underlying data.

Figure 14-7:      SwathPlot along X, Y, Z Direction for Ag (g/t) Validation

Source: SLR, 2023

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Figure 14-8:      Silver(g/t) Grade Section View Validation of Block Model – Section 22

Source: SLR, 2026

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Figure 14-9:      Silver(g/t) Grade Section View Validation of Block Model – Section DCAr0094

Source: SLR, 2026

14.14 Mineral Resource Statement

The QP has independently estimated the Project's Mineral Resources based on data collected by NPM. The Mineral Resource estimate and the underlying data comply with the guidelines provided in the CIM (2014) definitions under NI 43-101. Therefore, the QP is of the opinion that it is suitable for public reporting.

The Statement of Mineral Resources has been constrained by the topography and maximum optimized pit shell and reported using a 30 g/t AgEq cut-off grade. This cut-off value was calculated using the metal prices presented in Table 14-10 and the cost assumptions above.

Results of the independent Mineral Resources estimate for the Project are tabulated in the Statement of Mineral Resources within the three main zones: Upper Silver Zone, Middle Zinc Zone and Lower Gold Zone, shown in Table 14-11. Mineral Resources that are not mineral reserves do not have demonstrated economic viability.

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Table 14-11:      CarangasDeposit - Conceptual Pit* Constrained Mineral Resource as of March 31, 2026

Domain Category Density Tonnage AgEq Ag Au Pb Zn Cu
g/cm^3^ Mt Grade (g/t) Contained (Moz) Grade (g/t) Contained (Moz) Grade (g/t) Contained (Koz) Grade (%) Contained (Mlb) Grade (%) Contained (Mlb) Grade (%) Contained (Mlb)
Upper Silver Zone Indicated 2.09 121.6 70 272.4 44 173.9 0.06 220.0 0.34 923.3 0.65 1,729.6 0.01 34.91
Inferred 2.09 33.6 67 72.9 42 45.9 0.11 119.6 0.29 211.6 0.48 357.6 0.01 9.70
Middle Zinc Zone Indicated 2.29 38.7 41 51.3 12 15.1 0.06 68.8 0.37 314.1 0.81 688.0 0.02 12.96
Inferred 2.33 9.1 39 11.5 9 2.6 0.05 15.3 0.38 77.1 0.85 172.2 0.01 2.57
Lower Gold Zone Indicated 2.27 78.4 82 205.9 10 24.4 0.76 1,911.7 0.12 215.0 0.23 403.8 0.06 98.08
Inferred 2.33 11.0 80 28.3 10 3.6 0.70 248.5 0.15 36.2 0.34 82.4 0.05 11.78
Total Indicated 2.18 **238.**8 69 **529.**6 28 **213.**4 0.29 **2,200.**5 0.28 **1,452.**4 0.54 **2,821.**3 0.03 145.96
Inferred 2.18 **53.**8 65 **112.**7 30 **52.**1 0.22 **383.**4 0.27 **324.**9 0.52 **612.**3 0.02 24.05

* Notes:

1.       CIM (2014) definitions were followed for Mineral Resources.

2.       The QP for the Mineral Resource estimate, as defined by NI 43-101, is Anderson Candido, FAusIMM, Principal Geologist at SLR. The QP is independent of NPM and the Carangas Project.

3.       Mineral Resources that are not Mineral Reserves do not have demonstrated economic viability.

4.       The results are presented undiluted and on a dry in-situ basis and are considered to have reasonable prospects of economic viability.

5.       Mineral Resources are estimated at a cut-off grade of 30 g/t AgEq.

6.       AgEq formula is: AgEq g/t = Ag g/t + Au g/t * 80.49 + (Pb %*2204.6 /100 + Zn %*2866 /100 + Cu %* 8818.5 /100) / 1.318.

7.       Density varies from 2.09 in the upper zone to 2.33 in the lower zone.

8.       Mineral Resources are constrained by an optimized pit shell at a metal price of $41.00/oz Ag, $3,300.00/oz Au, $1.00/lb Pb, $1.30/lb Zn, $4.00/lb Cu. Recovery of 81.6% Ag, 93.4% Au, 73.4% Pb, 66.9% Zn, 38.7% Cu.

9.       1.85% of Mineral Resources are within the constrained pit but not within the mining concessions held by NPM.

10.     The drilling database was closed on June 1, 2023.

11.     The numbers may not add due to rounding.

Source: compiled by SLR, 2026

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14.15 Factors That May Affect theMineral Resource Estimate

The Mineral Resource for the Carangas Project is sensitive to several technical and non-technical factors that may influence both the tonnage and/or grade reported. The current geological model is based on drilling completed up to 2023, and no new drilling or geological updates have been incorporated since the 2023 Mineral Resource estimate. The geological model reflects the level of confidence supported by the available data but remains subject to refinement as additional geological information becomes available.

The geometry and orientation of the mineralized zones remain only partially constrained due to limited structural information and the absence of oriented core in most drill holes. This introduces potential uncertainty in the interpretation of domain boundaries, the continuity of high-grade zones, and the relationship between down-hole intercepts and true thickness. Variability inherent to epithermal systems including changes in breccia textures, alteration intensity, and metal zoning may also affect the distribution of grade within the model. In addition, the updated Mineral Resource relies on revised NSR assumptions reflecting updated metal prices, recoveries, and cost inputs; future changes to these assumptions may materially alter the constraining resource pit shell and therefore the reported Mineral Resource.

No Measured Mineral Resources have been defined for the Carangas Project at this stage. All Mineral Resources reported in this Technical Report fall within the Indicated and Inferred categories, reflecting the current level of drilling density, geological confidence, and the absence of sufficient data to support classification of any material at the Measured level. Future drilling, infill drilling, and improved geological and structural control will be required before any portion of the deposit can potentially meet the criteria for Measured classification under NI 43-101.

Other modifying factors may further influence the Mineral Resource Estimate as the Project advances. These include potential adjustments to metallurgical recoveries as test work evolves, geotechnical information that may affect pit slope design, hydrogeological conditions that may affect dewatering requirements, and the availability of surface land access for drilling and infrastructure. Environmental and social considerations, including permitting timelines and continued engagement with local communities, may also affect the ability to expand drilling coverage or convert Mineral Resources to Mineral Reserves in future technical studies. As more data are collected during subsequent project stages, these factors will need to be reassessed to refine and validate the Mineral Resource estimate.

The QP is not aware of any other factors, including environmental, permitting, legal, title, taxation, socio-economic, marketing, political, or other relevant factors that could materially affect the stated Mineral Resource estimate.

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15 Mineral Reserve Estimates
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This section is not relevant to this report.

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16 Mining Methods
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16.1 Introduction
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The deposit is amenable to open pit mining practices. Open pit mine designs, mine production schedules and mine capital and operating costs have been developed for the Carangas deposit at a PEA level of engineering. The Mineral Resources described in Section 14 form the basis of the mine planning, and includes both the Indicated and Inferred class resources.

Mine planning is based on conventional drill, blast, load, and haul open pit mining methods suited for the project location and local site requirements. The open pit activities are designed for approximately two years of construction followed by sixteen years of mine operations and, finally, three years of post-pit mining stockpile rehandle to the mill. The subset of Mineral Resources contained within the designed open pits are summarized in Table 16-1, with a US$13.50/t NSR cut-off and form the basis of the mine plan and production schedule.

Table 16-1:      PEAMine Plan Production Summary

Factor Value
PEA<br> Silver Zone Mill Feed 200.1<br> Mt
Silver<br> Zone Mill Feed NSR Grade US$42.2/t
Silver<br> Zone Mill Feed Ag Grade 36.2<br> g/t
Silver<br> Zone Mill Feed Pb Grade 0.34<br> %
Silver<br> Zone Mill Feed Zn Grade 0.64<br> %
PEA<br> Gold Mill Feed 51.4<br> Mt
Gold<br> Zone Mill Feed NSR Grade US$75.7/t
Gold<br> Zone Mill Feed Ag Grade 11.9<br> g/t
Gold<br> Zone Mill Feed Au Grade 0.74<br> g/t
Waste<br> Rock 363.4<br> Mt
Waste:<br> Resource Ratio 1.4

Notes:

1. The<br> PEA Mine Plan and Mill Feed estimates are a subset of the March 31, 2026, Mineral Resource<br> estimates and are based on open pit mine engineering and technical information developed<br> at a Scoping level for the Carangas deposit.
2. PEA<br> Mine Plan and Mill Feed estimates are mined tonnes and grade. The reference point is the<br> primary crusher.
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3. Net<br> Smelter Prices (NSP) and metallurgical recoveries define the cut-off grade. NSPs include<br> market price assumptions of $40.0/oz Ag, $1,984/t Pb, $2,646/t Zn, $3,200/oz Au. Various<br> smelter and refining terms, offsite costs, and a 5% royalty (6% for Ag) derive NSPs of $17.5/oz<br> Ag in Zn concentrate, $34.4/oz Ag in Pb concentrate, $34.8/oz Ag in doré, $3,028/oz<br> Au in doré, $1,070/t Pb, and $1,270/t Zn. Metallurgical recoveries of 6% Ag in Zn<br> concentrate, 81.6% Ag in Pb concentrate, 60.4% Ag in doré, 93.4% Au in doré,<br> 2.3% Pb in Zn concentrate, 73.4% Pb in Pb concentrate, 66.9% Zn in Zn concentrate, and 21.9%<br> Zn in Pb concentrate are applied.
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4. The<br> chosen cut-off grade covers total operating costs of $13.50/t, which exceeds estimated PEA<br> processing and G&A cost estimates.
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5. Estimates<br> have been rounded and may result in summation differences.
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Figure 16-1 shows the general arrangement for the PEA mine plan.

Figure 16-1:      MineOperations General Arrangements

Source: Moose Mountain, 2026

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Economic pit limits are determined using the Pseudoflow implementation of the Lerchs-Grossman algorithm. Selected pit limits are split up into five phases or pushbacks to target higher economic margin material earlier in the mine life. Upper benches of the open pit will be accessed via internal cut ramps on topography or via ramps left behind on phased pit walls. In-pit ramps will access material below the pit rim.

Pit designs are configured on 10 m bench heights, with a minimum of 8 m wide berms placed every two benches or double benching. Since no geotechnical test work or analysis has been completed on the bedrock, the applied bench face and inter-ramp angles, 67.5 degrees and 50 degrees, respectively, are scoping level assumptions based on the rock type and overall depth of the open pit.

Resources from the open pit will report to a ROM pad and primary crusher 0.5 km northeast of the pit rim. The mill will be fed with Silver Zone mineralized material from the pits at an average rate of 8.0 Mt/a (22 kt/d), increasing to 16.0 Mt/a (44 kt/d) in Year 6 of the Project. Between Years 9 and 16 of the Project, half the mill feed will be dedicated processing the gold zone mineralization.

Oxide resources will be stored in a stockpile 0.5 km east of the pit rim and rehandled to the crusher over the life of mine, blended with non-oxide mill feed. Non-oxide resources, mined in excess of mill feed targets, will be stored in a low-grade stockpile 1.5 km southeast of the ROM pad and process plant and east of the open pit. This stockpile is planned to be completely reclaimed to the mill at the end of the mine life.

Waste rock will be placed in a side hill waste rock storage facility (WRSF) 0 to 2.5 km north of the open pit, or used in the construction of haul roads and the dam portion of the tailings facility, which sits 3.0 km north of the open pit.

The waste rock from the open pit has not been tested or analyzed for potential acid generation (PAG). It is assumed that PAG quantities will be small enough to be blended with larger quantities of non-acid generating (NAG) waste rock for surface storage within the WRSFs.

Topsoil and overburden encountered at the top of the pits will be placed in a dedicated stockpile directly north of the open pit and kept salvageable for closure at the end of the mine life. These quantities have not been measured and the storage facility has not been designed for the PEA mine plan.

Contractor mining operations are planned utilizing a diesel-powered mining fleet. Cost estimates for mining are based on a contractor quotation for this Project, utilizing down-the-hole (DTH) drills for drilling, 0.20 kg/t target powder factor ANFO-based blasting, 9 m^3^ bucket size diesel hydraulic excavators for loading, and 90 t payload rigid-frame haul trucks for hauling, plus ancillary and service equipment to support the mining operations, including haul road and stockpile maintenance.

In-pit dewatering systems will be established for the pit. All surface water and precipitation in the open pit will be gravity drained or directed via pumps to ex-pit settling ponds directly outside the pit limits, where it will report to the broader project water management system.

Contractor cost estimates include investment in the mining mobile fleet and fixed facilities to maintain it.

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16.2 Key Design Criteria
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The following mine planning design inputs were used:

· Topography<br> is based on a LiDAR survey of the region in UTM Zone 20S coordinates. The topography surface<br> utilized for mine planning is not the same as used for Mineral Resources. Differences in<br> calculated pit contents are minimal and not material for this level of engineering.
· Resource<br> block models on 5 m spacing in all three dimensions, which contain whole block diluted metal<br> grades, bulk densities, and resource classifications.
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· Indicated<br> and Inferred class Mineral Resources are included in-pit optimizations and mill feed estimates.<br> Of the total planned mill feed, 80% is from Indicated class resources (160.8 Mt), and 20%<br> from Inferred class resources (39.3 Mt).
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· Scoping<br> level assumptions have been made for the pit configuration and overall pit slope angles based<br> on rock types and overall pit depth from surface. Open pit design inputs, including 67.5-degree<br> bench face angles, 50.0-degree interamp slope angles, and 20 m bench heights, are considered<br> reasonable for scoping level engineering on the project.
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· No<br> geographical restrictions have been applied to the open pit footprints. Stockpiles and rock<br> storage facilities are designed to avoid existing roads and streams.
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16.2.1 Net Smelter Prices, Net Smelter Return, and Cut-off Grade
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Mill feed and waste cut-off grade (COG) is based on the NSR in US$/t, which is determined using Net Smelter Prices (NSP). The NSR, net of offsite concentrate and smelter charges, including onsite mill recovery, is used as a cut-off item for break-even mill feed/waste selection.

The metal prices and smelter terms for the PEA economic assessment (see Section 22) are slightly different from the values described in this section and used for mine planning. Checks have been made by the QP to ensure that the PEA mine plan would not be materially altered by revising these inputs to the final PEA values.

NSP are used in place of metal market prices for mine planning to consider all offsite costs and determine revenue potential at the mine gate. The NSP calculation uses the inputs shown in Table 16-2.

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Table 16-2:      NetSmelter Price and Recoveries for Mine Planning

Item Value
Silver<br> (Ag) Price US$40/oz
Gold<br> (Au) Price US<br> $3,200/oz
Zinc<br> (Zn) Price US<br> $2,646/t
Lead<br> (Pb) Price US<br> $1,984/t
Ag<br> Deductions, Zn Con 93.3<br> g/t
Ag<br> Payable, Zn Con 70.0%
Ag<br> Deductions, Pb Con 50<br> g/t
Ag<br> Payable, Pb Con 96.5%
Ag<br> Payable, Dore 95.0%
Au<br> Payable, Dore 99.8%
Zn<br> Deductions, Con 8.0%
Zn<br> Payable, Con 85.0%
Pb<br> Deductions, Con 3.0%
Pb<br> Payable, Con 95.0%
Ag<br> Refining, in Con US<br> $0.50/oz
Ag<br> Refining, Dore US<br> $1.00/oz
Au<br> Refining, Dore US<br> $6.50/oz
Zn<br> Treatment US<br> $175/t Con
Pb<br> Treatment US<br> $100/t Con
Zn<br> Con Offsite Charges US<br> $35/t Con
Pb<br> Con Offsite Charges US<br> $120/t Con
Ag<br> Royalties 6.0%
Pb,<br> Zn, Au Royalties 5.0%
Ag<br> Process Recovery, Zn Con 6.0%
Ag<br> Process Recovery, Pb Con 81.6%
Ag<br> Process Recovery, Dore 60.4%
Au<br> Process Recovery, Dore 93.4%
Zn<br> Process Recovery, Zn Con 66.9%
Zn<br> Process Recovery, Pb Con 21.9%
Pb<br> Process Recovery, Zn Con 2.3%
Pb<br> Process Recovery, Pb Con 73.4%
Ag<br> NSP, Zn Con US<br> $0.56/g ($17.5/oz)
Ag<br> NSP, Pb Con US<br> $1.11/g ($34.4/oz)
Ag<br> NSP, Doré US<br> $1.12/g ($34.8/oz)
Au<br> NSP, Doré US<br> $97.34/g ($3,028/oz)
Zn<br> NSP $1,274/t
Pb<br> NSP $1,070/t
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The NSR calculation is shown in Equation 16.1.

Equation 16-1:NSR Calculation

Where:

Ag = silver grade (g/t)

Au = gold grade (g/t)

Zn = zinc grade (%)

Pb = lead grade (%)

RecAgZ = silver recovery in zinc concentrate (%)

RecAgP = silver recovery in lead concentrate (%)

RecAgD = silver recovery in doré (%)

RecAu = gold recovery (%)

RecZnZ = zinc recovery in zinc concentrate (%)

RecZnP = zinc recovery in lead concentrate (%)

RecPbZ = lead recovery in zinc concentrate (%)

RecPbP = lead recovery in lead concentrate (%)

NSPAgZ = net smelter price for silver in zinc concentrate ($/g)

NSPAgP = net smelter price for silver in lead concentrate ($/g)

NSPAgD = net smelter price for silver in doré ($/g)

NSPAu = net smelter price for gold ($/g)

NSPZn = net smelter price for zinc ($/t)

NSPPb = net smelter price for lead ($/t).

The breakeven economic cut-off grade (COG) is chosen as the NSR grade required to pay for processing costs, general and administration costs, and low-grade stockpile reclaim costs. The COG calculation uses the inputs shown in Table 16-3.

Table 16-3:      Cut-offGrade

Item Value
Process<br> and Tailings Costs $10.50/t
G&A<br> and Site Costs $1.50/t
Stockpile<br> Reclaim Costs $1.50/t
Chosen Project Economic Cut-off Grade $13.50/t

All resources above the chosen COG are either planned as direct mill feed from the pit or stockpiled and rehandled to the mill within the PEA mine plan.

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16.2.2 Mining Loss and Dilution
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The planning block model includes metal grades estimated on 5 m selective mining unit (SMU) size blocks. Additional loss and dilution calculations are based on the contacts between these 5 m SMU blocks.

There is an NSR value within each block of the planning model, and based on the breakeven economic cut-off grade, each block is identified as economic or uneconomic.

A procedure is run that examines each block, counts the number of contact edges between economic and uneconomic blocks, and stores this value, from 0 to 4, back into the planning model. Economic blocks with four uneconomic block contacts, or three block contacts and a grade below US$20/t NSR, are converted to waste since the increased costs to extract them selectively outweigh their economic benefit. Alternatively, uneconomic blocks with four block contacts, or three block contacts and grade above US$10/t NSR, are flagged and grouped as mill feed dilution for mine planning since the costs to separate them outweigh the impact to blending them in.

Figure 16-2 illustrates this concept, showing a plan view of the block model on the 3,800 masl elevation. The blocks are filled with colours based on their NSR grade. Orange outlines on the blocks signify if the block is treated as mill feed. All non-outlined blocks are treated as waste. Examples can be seen where isolated blocks are blended with surrounding blocks, either planned as loss or dilution depending on the original modelled grade.

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Figure 16-2:      MiningLoss and Dilution Application on 3800 masl bench

Source: Moose Mountain, 2024

This methodology introduces an additional 3% dilution, at the grade of the surrounding low-grade blocks, and a 97% mine recovery. This is over and above the dilution introduced in resource modelling to a 5 m SMU block size.

The PEA mine production plan (Section 16.8) is run using a 10 m bench elevation. It is recommended that future iterations of the mine planning block model use a 10 m SMU block size. Additional dilution introduced using a 10 m SMU block size is assumed to be negligible to the overall mill feed grades estimated in this PEA mine plan, based on the continuity within the 5 m SMU block model.

16.2.3 Production Rate Considerations

The PEA throughput target has been set at 8 Mt/a for the first five years of operations, increasing to 16 Mt/a in Year 6, setting the project life to 18.4 years. Half of the expanded throughput, or 8 Mt/a, will be allocated to processing gold zone mineralization between Years 9 and 16 of the Project.

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Several factors are considered when establishing an appropriate mining and processing rate. Key factors include:

· Resource<br> Size: Typically, a planned mine life is set at 12.5 to 20 years; beyond this, time-value<br> discounting shows an insignificant contribution to the NPV of the project at discount rates<br> of 8 % or higher.
· Capital<br> Payback: Capital investment typically is targeted at projects with a payback period of 2<br> to 5 years or shorter.
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· Operational<br> Constraints: Power, water, or supplies and services for support of operations can limit production.
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· Site<br> Delivery Constraints: Physical size and weight of equipment and shipping limits can determine<br> the maximum size of units that can be delivered to site.
--- ---
· Project<br> Financial Performance:
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o Generally,<br> economies of scale can be realized at higher production rates, which leads to reduced unit<br> operating costs. These are tempered to the above-mentioned physical and operational constraints<br> and flexibility issues.
--- ---
o Generally,<br> higher tonnage throughputs require more capital and the size of the project is reflected<br> in the initial investment. Economies of scale can still apply where some access and construction<br> issues have a high fixed component regardless of the project size.
--- ---
o Higher<br> production rates generally pay back fixed capital earlier and provide a higher rate of return<br> on capital, which improves project NPV.
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16.3 Pit Optimization
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The economic pit limits are determined using the Pseudoflow implementation of the Lerchs-Grossman algorithm. This algorithm uses the NSR grades and bulk density for each block of the 3D block model and evaluates the costs and revenues of the blocks within potential pit shells. The routine uses input economic and engineering parameters and expands downwards and outwards until the last increment is at break-even economics.

Additional cases are included in the analysis to evaluate the sensitivities of open pit mined resources to waste mining ratio and high-grade/low-grade areas of the deposits. In this study, the various cases or pit shells are generated by varying the input metal prices and comparing the resultant waste and mill feed tonnages and metal grades for each pit shell.

By varying the economic parameters while keeping inputs for metallurgical recoveries and pit slopes constant, various generated pit cases are evaluated to determine where incremental pit shells produce marginal or negative economic returns. This drop-off is due to increasing waste mining ratios, decreasing metal grades, increased mining costs associated with the larger or deeper pit shells, and the time value of discounting revenues more than costs. The economic margins from the expanded cases are evaluated on a relative basis to provide payback on capital and produce a return for the project. At some point, further expansion does not provide significant added value. A pit limit can then be chosen that has a suitable economic return for the deposit.

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An undiscounted cashflow (UCF) is generated for each pit shell based on the shell contents and the economic parameters listed in Table 16-4. The UCF for each case is compared to reinforce the selected point at which increased pit expansions do not increase the project value. Note that the economics are only applied for comparative purposes to assist in the selection of an optimum pit shell for further mine planning; they do not reflect the actual financial results of the mine plan.

The chosen pit shell is then used as the basis for more detailed design and economic modelling.

The metal prices are varied from 10% to 150% of the market prices listed in Table 16-2.

Table 16-4:      OperatingCost Inputs for Pseudoflow Pit Shells

Item Value
Pit<br> Rim Mining Cost (mill feed and waste) $2.50/t<br> mined
Incremental<br> Haulage Cost $0.01/t<br> mined per 5 m drop below 3,910 masl
Process<br> and Tailings Cost $10.50/t<br> milled
General<br> and Administrative Cost $1.50/t<br> milled
16.3.1 Ultimate Pit Limits
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Figure 16-3 shows the contents of the generated Pseudoflow pit shells for the Carangas deposit. An inflection point can be seen in the curve of cumulative resources and undiscounted cash flow (UCF) by pit case. This point indicates Price Factor (PF) Case 0.70 as a point at which larger pit shells will not produce material increases to project value.

The pit shell generated from 0.70 PF case is selected as the ultimate pit limits for Carangas and is used for further mine planning as a target for detailed open pit designs with berms and ramps.

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Figure 16-3:      PseudoflowPit Shell Resource Contents by Case

Source: MMTS, 2026

16.4 Pit Designs
16.4.1 Open Pit Phases
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Ultimate pit limits are generally split up into phases or pushbacks to target higher economic margin material earlier in the mine life. Minimum pushback distances of 50 m are honoured. The pit is split into five phases, with the higher-grade, lower strip ratio early pit phases mined ahead of lower grade, higher strip ratio pushbacks to the ultimate pit limit. The final two phases will target the gold zone mineralization. Targets for the initial and interim pit phases use Case PF 0.25, Case PF 0.35, Case PF 0.39, and Case PF 0.45 of the optimization runs described in Section 16.3.1. In Year 1, the Carangas River will be relocated to the east of the open pit to accommodate the expansion of the open pit over the LOM.

16.4.2 In Pit Haul Roads

In-pit haul roads are designed 28 m wide to facilitate two-way travel for 140 t payload rigid frame haul trucks. Haul road grades are limited to a maximum of 10%. Access ramps are not designed for the last bench (10 m) of the pit bottom, on the assumption that the bottom ramp segment will be removed using some form of retreat mining. The next bottom bench (10 m) of the pit uses one-way haul roads of 21 m width and 12% grade since bench volumes and traffic flow are reduced.

Benches above the pit rim exit can be accessed by external roads built on the original hill side slopes reducing the need for in-pit haul ramps in the final pit wall above the pit exit.

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16.4.3 Open Pit Contents

Contents of the designed open pits are presented in Table 16-5. The contents for each designed pit phase are presented graphically in Figure 16-4.

Table 16-5:     DesignedOpen Pit Contents

Pit Phase Pit Name Mill Feed (Mt) NSR Grade ($/t) Ag Grade (g/t) Au Grade (g/t) Pb Grade (%) Zn Grade (%) Waste (Mt) W:R Ratio (t/t)
Starter Pit P621 39.4 64.4 58 0.01 0.42 0.72 26.7 0.7
SW Pushback P622i 29.9 48.8 40 0.04 0.37 0.83 28.0 0.9
NE Pushback P623i 40.0 42.2 37 0.03 0.33 0.57 24.6 0.6
North Pushback P624i 75.9 38.5 25 0.21 0.29 0.51 141.3 1.9
Final Pushback P625i 66.4 56.3 15 0.47 0.22 0.43 142.9 2.2
Total P625 251.5 49.1 31 0.20 0.31 0.57 363.4 1.4

Notes :

1. The PEA Mine Plan and Mill Feed estimates are a subset of the March 31, 2026, Mineral Resource estimates<br>and are based on open pit mine engineering and technical information developed at a Scoping level for the Carangas deposit.
2. PEA Mine Plan and Mill Feed estimates are mined tonnes and grade. The reference point is the primary crusher.<br>Mill Feed tonnages and grades include open pit mining method modifying factors, such as dilution and recovery.
--- ---
3. Net Smelter Prices (NSP) and metallurgical recoveries define the cut-off grade. NSPs include market price<br>assumptions of $40.0/oz Ag, $1,984/t Pb, $2,646/t Zn, $3,200/oz Au. Various smelter and refining terms, offsite costs, and a 5% royalty<br>(6% for Ag) derive NSPs of $17.5/oz Ag in Zn concentrate, $34.4/oz Ag in Pb concentrate, $34.8/oz Ag in doré, $3,028/oz Au in doré,<br>$1,070/t Pb, and $1,270/t Zn. Metallurgical recoveries of 6% Ag in Zn concentrate, 81.6% Ag in Pb concentrate, 60.4% Ag in doré,<br>93.4% Au in doré, 2.3% Pb in Zn concentrate, 73.4% Pb in Pb concentrate, 66.9% Zn in Zn concentrate, and 21.9% Zn in Pb concentrate<br>are applied.
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4. The chosen cut-off grade covers total operating costs of $13.50/t, which exceeds estimated PEA processing<br>and G&A cost estimates.
--- ---
5. Of the total mill feed, 4.6 Mt at 39.3 g/t Ag, 0.26% Pb, 0.55% Zn are located outside of NPM concessions.
--- ---
6. Estimates have been rounded and may result in summation differences.
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Figure 16-4:      DesignedOpen Pit Contents

Source: Moose Mountain, 2026

16.4.4 Open Pit Designs

The pit designs are shown in Figure 16-5 (final pit phase) and Figure 16-6. Original topography contour polylines are shown on 5 m vertical intervals. Sections through the deposit showing the resource model grades are illustrated in Figure 16-7 and Figure 16-8.

16.4.4.1 Starter Phase:

This phase targets the higher-grade, lower-strip ratio portion of the deposit outlined by the Case PF 0.25 pit shell described in Section 16.3.1. It contains enough mill feed mineralization for the first five years of 8 Mt per year operations. The upper benches of this phase will be accessed via in-pit cut ramps up to 4,070 masl in the northwest and 3,990 masl in the east, developed during the project's construction period. Pit ramps are left behind in the high wall for access to future west high wall pushbacks. These ramps run from the 4,020 masl elevation in the north down to the pit exit at the 3,910 masl elevation in the northeast. In-pit ramping is also incorporated from the pit exit, running counterclockwise down to a pit bottom in the center of the pit at 3,800 masl and switchbacking at the 3,840 masl and 3,810 masl elevations to a separate pit bottom in the southwest at 3,760 masl. A saddle between these two pit bottoms is left behind at 3,840 masl.

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16.4.4.2 SW Pushback Phase:
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This phase targets deeper, lower grade mineralization below the starter pit in the southwest, outlined by the Case PF 0.35 pit shell described in Section 16.3.1. It contains about two years of expansion mill feed mineralization at 16 Mt per year. The pit highwall is pushed to the final pit limits in the southwest. Benches between 3,975 masl and the pit exit at 3,910 masl will utilize in-pit ramps left behind in the starter pit walls. In-pit ramping is also incorporated from the pit exit, running clockwise down to the pit bottom on the 3,680 masl elevation in the southwest. A separate pit bottom is established in the center of the pit, between the 3,780 masl and 3,750 masl elevations.

16.4.4.3 NE Pushback Phase:

This phase pushes the pit out to the north and east, targeting deeper, lower grade mineralization below the starter pit in the north, outlined by the Case PF 0.39 pit shell described in Section 16.3.1. It contains about 2.5 years of expansion mill feed mineralization. The upper benches of the northwest highwall will be accessed via in-pit cut ramps developed up to 4,070 masl and down to the pit exit at 3,910 masl. Ramping is left behind on this highwall between 3,980 masl and 3,910 masl to access upper benches of future pit pushbacks. Ramping is also left behind on the east highwall, between 4,000 masl and 3,910 masl to access upper benches of future pit pushbacks. In-pit ramping is incorporated from the pit exit, running clockwise down to the pit bottom at 3,740 masl, switching back to counterclockwise at 3,870 masl.

16.4.4.4 North Pushback Phase:

This phase targets the initial portions of the deeper gold zone mineralization, outlined by the Case PF 0.45 pit shell described in Section 16.3.1. It contains about four years of expansion mill feed mineralization, as well as two years of gold zone mill feed at 8 Mt per year. This phase pushes the highwall to the ultimate limits in the west, north and east. The upper benches of the west and east highwall will be accessed via in-pit cut ramps developed in previous phases. In-pit ramping is incorporated from the pit exit, at 3,910 masl, running clockwise down to the pit bottom at 3,560 masl.

16.4.4.5 Final Pushback Phase:

This final pit phase targets the economic open pit limits of the resource, extending the highwall to ultimate limits in the south and southeast. It contains about two years of expansion mill feed mineralization, as well as 4.5 years of gold zone mill feed at 8 Mt/a. In-pit ramping is incorporated from the pit exit, at 3,910 masl, running clockwise down to a pit bottom at 3,380 masl, switchback at 3,790 masl and 3,710 masl.

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Figure 16-5:       UltimatePit Design, P621

Source: Moose Mountain, 2026

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Figure 16-6:       PhasedPit Designs

Source: Moose Mountain, 2026

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Figure 16-7:       PitDesigns, EW Section View, 7905350N

Source: Moose Mountain, 2026

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Figure 16-8:       PitDesigns, NS Section View, 539150E

Source: Moose Mountain, 2026

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16.5 Low Grade and Oxide Stockpile Design
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When resources are mined from the pit, they will either be delivered to the primary crusher, the ROM stockpile located next to the crusher, or the low grade and oxide stockpiles.

The primary crusher and ROM stockpiles are located 0.5 km northeast of the pit limits.

Cut-off grade optimization on the mine production schedule sends resources between US$13.50/t NSR and US$35/t NSR to a low-grade stockpile located 1.5 km east of the ROM pad and open pit. These stockpiled resources are planned to be re-handled back to the crusher once the open pit is exhausted.

Oxide resources mined from the open pit will be blended with non-oxide resources as part of the overall mill feed plan. When this blending cannot be done at the ROM pad/primary crusher, the oxide resources will be stored in a stockpile directly east of the pit rim and rehandled to the crusher over the LOM. Target for blending is 13% oxide in the silver zone mill feed.

Preliminary designs for these facilities are completed assuming:

· bottom-up construction and top-down reclamation
· 18-degree overall slopes (3:1)
--- ---
· storage density of 2.0 t/m^3^
--- ---
· maximum height from topography to crest of 65 m on the low-grade stockpile and 55 m on the oxide stockpile.
--- ---

The low grade and oxide stockpiles are shown in the mine area general arrangement drawing in Figure 16-1.

16.6 Waste Rock Storage Facility Design

Waste rock mined from the open pit will be placed in the waste rock storage facilities (WRSF), designed with an 18-degree overall slope (3:1) and a storage density of 2.0 t/m^3^, or placed as construction fill for haul roads or the tailings facility dam.

The WRSF runs for 2.5 km between the open pit and north towards the tailings facility. The facility sits on a hillside and will be constructed from a combination of hillside dumping and bottom-up lift construction. Waste rock from the upper benches of the open pit will be sent along topography to the upper lifts of this facility, minimizing haulage requirements for this waste rock.

The waste rock from the open pit has not been tested or analyzed for potential acid generation (PAG). It is assumed that the majority of the waste rock is net acid neutralizing, and there has been no consideration for the segregation of different rock types in the planned storage facilities. Further test work and analysis is recommended to better classify waste materials according to acid-generating potential and to confirm that a blending strategy is the preferred method for handling and storing any potentially acid-generating waste rock.

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Backfilling of the open pit was examined as an opportunity, but space is too limited as the pit continually expands deeper in all directions until the final pit phase.

Topsoil and overburden encountered at the top of the pits will be placed in dedicated stockpiles directly north or south of the open pit and kept salvageable for closure at the end of the mine life. Quantities of this material will be minimal, but an estimate has not been made for this PEA mine plan, and potential storage facilities have not been designed.

The waste storage facility is shown in the mine area general arrangement in Figure 16-1.

16.7 Ex-Pit Haul Roads

Mine haul roads external to the open pits are planned to connect the open pit to scheduled destinations for hauling resource and waste materials.

Haul road designs are limited to 34 m wide outlines along preferred routes. 3D cut and fill designs have not been completed for this PEA mine plan. Routes have been chosen to maximize fill design potential, with fill rock planned to be sourced from the open pit during the construction period of the project. Estimated mining costs during the construction period will cover the construction costs of these ex-pit haul roads.

The ex-pit haul layouts in the mine area general arrangement in Figure 16-1.

16.8 Mine Production Schedule

Mill feed requirements by scheduled period, mine operating considerations, product prices, recoveries, destination capacities, equipment performance, haul cycle times and operating costs are used to determine the optimal production schedule from the phased pit contents.

The overall production schedule is included in Table 16-6, with mill feed tonnes and grade illustrated in Figure 16-9, and overall mine production tonnage and waste mining ratio illustrated in Figure 16-10, and phases mined illustrated in Figure 16-11.

The production schedule is based on the following parameters:

· The Mineral Resource and associated waste material quantities are split by pit phase and bench quantities.
· The operations are scheduled on annual periods.
--- ---
· A mill feed rate of 8,000 kt/a (22 kt/d) is targeted for the first 5 years of operations.
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o A first-year mill throughput ramp-up target of 7,250 kt is assumed.
--- ---
o Expansion of silver zone mill feed to 16,000 kt/a (44 kt/d) in Year 6 of operations.
--- ---
o Conversion of 8,000 kt/a to gold zone processing, starting halfway through Year 9 of operations and lasting<br>until the start of Year 16 of operations. The remaining 8,000 kt/a will continue to process silver zone mineralization.
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o A significant quantity of oxide resources, as well as some non-oxide resources, are planned to be stockpiled<br>well in advance of the mill ramp-up period.
--- ---
o Oxide mill feed of 13% of overall silver zone mill feed.
--- ---
· Within a given pit phase, each bench is fully mined before progressing to the next bench.
--- ---
· Pit phases are mined in sequence, where the later pit phases do not mine below former pit phases.
--- ---
· Pit phase vertical progression in the mineralized area is limited to no more than 90 m each year, or nine<br>benches; the average annual phase progression is 50 m.
--- ---
· Pre-stripping done in the construction period, Years -2 and -1, is done to open the pits sufficiently<br>to supply non-oxide mill feed at the target throughput rate in Year 1 of the Project.
--- ---
o This also provides sufficient construction waste rock for the mine haul roads and the starter tailings<br>facility dam.
--- ---
· Resource tonnes released in excess of the mill capacity are stockpiled, including those mined in the construction<br>phase.
--- ---
· Shovel and haul truck operating hour estimates are run as part of the mine schedule. Haul cycle times<br>are simulated from all pit benches to all destinations. Total pit production is balanced on calculated hauler operating hour requirements.<br>This strategy is used to avoid large spikes and dips in the number of haulers in the LOM schedule but leads to some variations in total<br>tonnes mined in each period. Cycle time simulations should be refined in future engineering studies.
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Figure 16-9:       AnnualMill Feed Tonnes and Grade

Source: Moose Mountain, 2026

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Figure 16-10:       AnnualMaterial Mined and Waste Mining Ratio

Source: Moose Mountain, 2026

Figure 16-11:       PitPhases Mined

Source: Moose Mountain, 2026

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Table 16-6:       MineProduction Schedule

Mine Production: Year LOM Y-2 Y-1 Y01 Y02 Y03 Y04 Y05 Y06 Y07 Y08 Y09 Y10 Y11 Y12 Y13 Y14 Y15 Y16 Y17 Y18 Y19
Silver Zone Mill Feed Mt 200.1 0.0 0.0 7.2 8.0 8.0 8.0 8.0 16.0 16.0 16.0 11.5 8.0 8.0 8.0 8.0 8.0 8.0 15.0 16.0 16.0 6.3
NSR $/t 42 0 0 58 67 65 67 72 54 59 54 49 35 34 28 29 29 29 27 20 20 17
silver (Ag) g/t 36 0 0 50 59 59 61 65 48 54 51 46 28 28 22 23 23 23 21 13 13 9
Lead (Pb) % 0.34 0.00 0.00 0.50 0.51 0.42 0.37 0.49 0.36 0.35 0.27 0.26 0.35 0.33 0.30 0.28 0.28 0.28 0.29 0.33 0.33 0.31
Zinc (Zn) % 0.64 0.00 0.00 0.87 0.91 0.79 0.73 0.94 0.71 0.59 0.43 0.39 0.63 0.60 0.64 0.59 0.59 0.59 0.60 0.65 0.65 0.70
Oxide Mill Feed Mt 25.6 0.0 0.0 1.0 1.1 1.1 1.1 1.1 2.1 2.1 2.1 1.5 1.1 1.1 1.1 1.1 1.1 1.1 2.1 2.1 2.1 0.0
Transition Mill Feed Mt 6.0 0.0 0.0 1.4 0.8 0.7 0.1 0.3 0.3 0.0 0.0 0.0 0.0 0.1 0.2 0.3 0.3 0.3 0.4 0.2 0.2 0.1
Fresh Mill Feed Mt 168.5 0.0 0.0 4.9 6.1 6.2 6.8 6.6 13.6 13.9 13.9 10.0 6.9 6.8 6.7 6.7 6.7 6.7 12.5 13.7 13.7 6.2
Gold Zone Mill Feed Mt 51.4 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 3.5 8.0 8.0 8.0 8.0 8.0 7.0 1.0 0.0 0.0 0.0
NSR $/t 76 0 0 0 0 0 0 0 0 0 0 55 48 64 59 77 96 112 175 0 0 0
silver (Ag) g/t 12 0 0 0 0 0 0 0 0 0 0 13 11 13 10 16 13 9 8 0 0 0
Gold (Au) % 0.74 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.00 0.51 0.45 0.61 0.58 0.73 0.96 1.16 1.86 0.00 0.00 0.00
Resource Mined from Pit Mt 251.5 1.6 5.9 11.7 12.8 14.1 7.6 12.6 26.2 27.5 29.6 27.0 20.7 17.5 12.5 8.2 8.1 7.0 1.0 0.0 0.0 0.0
NSR $/t 49 42 44 56 54 52 66 55 42 44 39 34 34 43 44 76 95 111 175 0 0 0
Mined Directly to Mill Mt 146.6 0.0 0.0 6.3 7.0 7.0 6.9 7.0 13.9 13.9 13.9 12.4 13.9 11.7 8.6 8.0 8.1 7.0 1.0 0.0 0.0 0.0
NSR $/t 65 0 0 60 70 68 70 77 56 61 56 51 42 55 57 77 95 111 175 0 0 0
LGSP Balance Mt - 0.0 0.0 0.0 1.9 5.0 5.5 9.0 19.3 32.4 48.0 61.8 67.5 70.1 67.6 60.8 53.9 47.0 34.1 20.2 6.3 0.0
NSR $/t - 40 38 38 23 25 24 25 24 25 25 24 23 22 21 21 20 19 17 17 17 0
Oxide SP Balance Mt - 1.6 7.5 12.0 14.9 18.0 17.0 18.1 17.8 16.2 14.1 12.6 11.6 10.5 9.5 8.4 7.4 6.3 4.2 2.1 0.0 0.0
NSR $/t - 42 43 47 45 45 45 43 43 42 42 42 42 42 42 42 42 42 42 42 42 0
Waste Mined Mt 363.4 3.4 8.2 16.8 13.4 12.4 16.6 12.4 58.8 50.9 38.3 32.4 30.3 27.5 27.6 9.3 4.1 1.0 0.0 0.0 0.0 0.0
Waste to WRSF Mt 173.9 0.0 0.0 1.8 0.4 0.4 2.6 0.4 44.8 36.9 24.3 18.4 16.3 13.5 13.6 0.3 0.1 0.0 0.0 0.0 0.0 0.0
Waste to Haul Roads Mt 2.0 2.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0 0.0
Waste to Tailings Dam Mt 187.5 1.4 8.2 15.0 13.0 12.0 14.0 12.0 14.0 14.0 14.0 14.0 14.0 14.0 14.0 9.0 4.0 1.0 0.0 0.0 0.0 0.0
Waste/Mill Feed Mined - 1.4 0.0 0.0 2.3 1.7 1.5 2.1 1.6 3.7 3.2 2.4 2.2 1.9 1.7 1.7 0.6 0.3 0.1 0.0 0.0 0.0 0.0
Cumulative Ratio - - 0.0 0.0 3.9 2.7 2.3 2.3 2.1 2.6 2.7 2.6 2.6 2.5 2.4 2.3 2.2 2.0 1.8 1.7 1.6 1.5 1.4
Total Material Mined Mt 615.0 5.0 14.1 28.6 26.2 26.5 24.2 25.0 85.0 78.4 67.9 59.4 51.0 45.0 40.1 17.5 12.2 8.0 1.0 0.0 0.0 0.0
Total Material Moved Mt 719.9 5.0 14.1 29.5 27.2 27.5 25.3 26.0 87.1 80.5 70.0 62.0 53.1 49.3 47.4 25.5 20.1 16.0 16.0 16.0 16.0 6.3
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16.8.1 End of Period Figures
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Figure 16-12 to Figure 16-17 shows the progression of the open pit and stockpiles in Years -1, 1, 3, 5, 9 and end of mine life.

Figure 16-12:       Endof Period Mine Production Schedule, Year -1

Source: Moose Mountain, 2026

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Figure 16-13:       Endof Period Mine Production Schedule, Year 1

Source: Moose Mountain, 2026

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Figure 16-14:       Endof Period Production Schedule, Year 3

Source: Moose Mountain, 2026

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Figure 16-15:       Endof Period Production Schedule, Year 5

Source: Moose Mountain, 2026

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Figure 16-16:       Endof Period Production Schedule, Year 9

Source: Moose Mountain, 2026

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Figure 16-17:       Endof Period Production Schedule, Year 19

Source: Moose Mountain, 2026

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16.9 Mine Operations
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Contractor-operated and managed open pit mine operations are planned to be typical of similar terrain operations in South America.

An owner’s management and technical team is planned to provide the mining contractor with geology, engineering, and surveying support and management guidance.

In situ rock will be drilled and blasted to create suitable fragmentation for efficient loading and hauling of both resource and waste rock. Blasthole sampling and assaying will provide bench scale grade control direction. Loading will be completed with a hydraulic excavator, with dig lines between resources and waste directed by the grade control program. Resource and waste rock will be hauled out of the pit and to scheduled destinations with off-highway rigid-frame haul trucks.

In-pit dewatering systems will be established for the pit. All surface water and precipitation in the open pit will be gravity drained or directed via submersible pumps to ex-pit settling ponds directly outside the pit limits, where it will report to the wider project water management system.

Other mine pit services will include:

· haul road maintenance
· pit floor and ramp maintenance
--- ---
· mobile fuel and lube services
--- ---
· secondary blasting and rock-breaking
--- ---
· lighting
--- ---
· transporting personnel and operating supplies
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· mobile maintenance services.
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Mining operations are based on 365 operating days per year.

The contractor bids for this work envision utilizing a diesel-powered mining fleet. The contractor would bring their own fleet, which includes DTH drills for production drilling, 0.20 kg/t target powder factor ANFO-based blasting, 6 m^3^ bucket size diesel hydraulic excavators for loading, and 90 t payload rigid-frame haul trucks for hauling, plus ancillary and service equipment to support the mining operations. The contractor will also maintain their mining fleet in the field and within contractor-supplied fixed maintenance facilities.

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16.10 Risks
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The project is at a scoping engineering level. Limited geotechnical, hydrogeological, and geochemical information and data have been collected across the Project. Further fieldwork, lab work, and modelling are required to advance the engineering to the next stage of Pre-feasibility or Feasibility. It can be anticipated that further field drilling and advancement of the project engineering will materially alter the existing mine plan, reducing the plan’s risk and identifying and exploiting the potential opportunities that arise.

Risks to the PEA defined mill feed quantities, metal grades, associated waste rock quantities and the estimated costs to exploit include changes to the following factors and assumptions:

· Metal prices.
o Significant (>50%) decreases in metal prices may increase the economic cut-off grade or reduce the<br>size of the selected open pit limits, with either outcome reducing the size of the resource base to include into the mine plan.
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· Interpretations of mineralization geometry and continuity in mineralization zones.
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o Decreases in the resource base could significantly alter the mine plan.
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· Geotechnical and hydrogeological assumptions.
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o Geotechnical sampling, test work, and analysis may show a required shallowing of pit slope angles, which<br>would likely increase the overall LOM stripping ratio to access the resource.
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o Hydrogeological sampling, test work, and analysis may identify the need for a more onerous (costly) pit<br>water management and pit slope depressurization solution.
--- ---
· Geochemical assumptions for mined resources and waste materials.
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o Geochemical sampling, test work, and analysis, specifically in the open pit waste rock, may identify a<br>more onerous (costly) PAG management solution.
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· Ability of the mining and milling operation to meet the annual production rate and anticipated grade control<br>standards and recoveries.
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o Reduced selectivity with the mining fleet, reduced mining or milling recoveries, or increased mining dilution<br>would result in an increased cost of achieving the planned PEA metal production.
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· The ability of the milling operation to meet the annual production rate and recoveries.
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· Operating cost assumptions and cost creep.
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o Mining cost assumptions are based on a quote from a contractor with extensive in-country operating experience.<br>However, further detailing of contractor costs may also lead to different operational cost estimates.
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· Ability to meet and maintain future land tenure, permitting, and environmental license conditions, and<br>the ability to maintain the social license to develop and operate.
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· Ability to access capital for project financing.
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17 Recovery Methods
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17.1 Overview
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The proposed processing strategy for the Carangas Project is based on a staged development of processing capacity over the LOM. During the initial operating period (Phase 1), Years 1 to 5, mineralised material will be treated through a single silver-lead-zinc flotation process plant with a nominal design throughput of 22,000 t/d to produce a silver-lead concentrate and a zinc concentrate. Processing capacity is planned to increase during the expansion period (Phase 2), Years 6 to 8, to a nominal 44,000 t/d through the installation of a twinned 22,000 t/d process plant, duplicating the principal unit operations and associated infrastructure required for silver-lead and zinc concentrate production. In Year 9, one of the two process plants is planned to be converted (comminution and rougher flotation to be used) for the treatment of gold-bearing material and will operate as a dedicated 22,000 t/d gold processing facility after adding dedicated gold recovery equipment including cyanide leaching, counter-current decantation thickeners, Merrill Crowe and refining (Phase 3). One of the silver-lead-zinc plants will continue to operate as a 22,000 t/d concentrator. This dual-processing configuration is expected to continue until Year 16, with the gold plant converted back to a concentrator circuit for the remainder of the LOM. The phased approach provides flexibility to align processing capacity and flowsheet configuration with the mine production schedule and evolving feed characteristics.

The process flowsheet for the project is based on preliminary metallurgical laboratory testing, as discussed in Section 13, and preliminary economic modelling. The selected unit operations are conventional technologies commonly applied in silver, lead, zinc, and gold processing plants of similar throughput, operating 24 hours per day and 365 days per year. The proposed flowsheet includes the following process areas and unit operations:

· Crushing – primary crushing followed by stockpiling
· Grinding – semi-autogenous grinding (SAG) mill, followed by ball milling with cyclone classification
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· Silver-lead flotation – rougher, interstage thickening, concentrate regrind and three-stage cleaning
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· Silver-lead concentrate handling – concentrate thickening and filtration
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· Zinc flotation – rougher, concentrate regrind and three-stage cleaning
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· Zinc concentrate handling – concentrate thickening and filtration
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· Gold flotation – rougher flotation and thickening of rougher tailings
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· Gold leaching – pre-leach thickening and cyanide leaching followed by counter-current decantation<br>(CCD)
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· Gold recovery –zinc precipitation of the clarified pregnant solution and smelting to produce doré
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· Cyanide destruction and tailings management – tailings cyanide destruction, followed by tailings<br>thickening.
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The feed grades and recoveries are summarized in Table 17-1 and Table 17-2.

Table 17-1:       PlantFeed Grades

Description Units Years 1 to 8 Years 9 to 16 LOM
Silver Zone
Silver feed grade g/t 55.92 26.66 36.22
Lead feed grade % 0.41 0.30 0.34
Zinc feed grade % 0.74 0.58 0.64
Gold Zone
Silver feed grade g/t - 11.92 11.92
Gold feed grade g/t - 0.74 0.74

Table 17-2:       PlantRecoveries

Description Units Years 1 to 8 Years 9 to 16 LOM
Silver-Lead<br> Concentrate
Silver<br> recovery % 80.5 77.8 79.4
Lead<br> recovery % 71.6 70.5 71.5
Zinc<br> recovery % 21.9 21.9 21.9
Zinc<br> Concentrate
Silver<br> recovery % 4.7 4.3 4.5
Lead<br> recovery % 2.3 2.3 2.3
Zinc<br> recovery % 61.7 61.8 62.1
Gold<br> Doré
Silver<br> recovery % - 60.0 60.0
Gold<br> recovery % - 93.0 93.0
17.2 Process Flowsheet
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The process flowsheet for the concentrate plant is shown in Figure 17-1, while the process flowsheet for the gold plant is shown in Figure 17-2.

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Figure 17-1:       ConcentrateProcess Plant Flowsheet

Source: Ausenco, 2026

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Figure 17-2:       GoldProcess Plant Flowsheet

Source: Ausenco, 2026

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17.3 Plant Design
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The main design parameters for the concentrate plant are presented in Table 17-3, while the main design parameters for the gold plant are presented in Table 17-4. The major process equipment sizing is presented in Table 17-5.

Table 17-3:       ConcentratePlant Process Design Criteria

Description Units Value
Design Basis
Annual<br> throughput (dry) Mt/a 8.0
Daily<br> average throughput (dry) t/d 22,000
Operating Availability
Primary<br> crushing % 75
Grinding<br> and flotation % 91.3
Concentrate<br> filtration % 89
Design Feed Grades
Silver<br> (Ag) g/t 47.0
Lead<br> (Pb) % 0.37
Zinc<br> (Zn) % 0.67
Design Concentrate Grades
Silver-Lead<br> concentrate silver grade, minimum g/t 2000
Silver-Lead<br> concentrate lead grade % 24
Zinc<br> concentrate zinc grade, minimum % 45.8
Design Recoveries
Silver<br> recovery to Ag/Pb concentrate % 79.4
Lead<br> recovery to Ag/Pb concentrate % 71.5
Silver<br> recovery to Ag/Zn concentrate % 4.5
Zinc<br> recovery to Ag/Zn concentrate % 62.1
Run-of-Mine (ROM) Characteristics
Solids<br> specific gravity - 2.74
Moisture<br> content %<br> (w/w) 5
Comminution Characteristics
JK<br> SMC test parameters (A x b), design – client provided - 48
Bond<br> ball mill work index (BWi), 75th percentile kWh/t 12.8
Abrasion<br> index (Ai), 75th percentile g 0.054
Crushing
Crushing<br> circuit product size, P80 mm 125
Stockpile<br> live capacity at reclaim rate h 16
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Description Units Value
--- --- ---
Grinding
Grinding<br> circuit product size, P80 μm 75
Ag/Pb Flotation Circuit
Residence<br> time, scale-up factor - 2.5
Rougher<br> residence time, from laboratory test min 14
Rougher<br> concentrate regrind product size, P80 μm 20
Regrind<br> specific energy kWh/t 9.6
Cleaner<br> 1 residence time, from laboratory test min 5.0
Cleaner<br> 2 residence time, from laboratory test min 4.0
Cleaner<br> 3 residence time, from laboratory test min 3.0
Zn Flotation Circuit
Residence<br> time, scale-up factor - 2.5
Rougher<br> residence time, from laboratory test min 6
Rougher<br> concentrate regrind product size, P80 μm 25
Regrind<br> specific energy kWh/t 10.6
Cleaner<br> 1 residence time, from laboratory test min 2.0
Cleaner<br> 2 residence time, from laboratory test min 2.0
Cleaner<br> 3 residence time, from laboratory test min 2.0
Ag/Pb Concentrate Handling
Concentrate<br> thickener unit area settling rate t/m^2^/h 0.25
Concentrate<br> thickener underflow pulp density %<br> solids (w/w) 60
Cycle<br> time min 13.5
Filter<br> cake moisture %<br> (w/w) 10
Zn Concentrate Handling
Concentrate<br> thickener unit area settling rate t/m^2^/h 0.15
Concentrate<br> thickener underflow pulp density %<br> solids (w/w) 60
Cycle<br> time min 13.2
Filter<br> cake moisture %<br> (w/w) 10
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Table 17-4:       GoldPlant Process Design Criteria

Description Units Value
Design Basis
Annual<br> throughput (dry) Mt/a 8.0
Daily<br> average throughput (dry) t/d 22,000
Operating Availability
Primary<br> crushing % 75
Grinding<br> and flotation % 91.3
Design Feed Grades
Gold<br> (Au) g/t 0.8
Silver<br> (Ag) g/t 15.6
Flotation Recovery, at LOM average grade
Gold<br> recovery to doré % 93.0
Silver<br> recovery to doré % 60.0
Run-of-Mine (ROM) Characteristics
Solids<br> specific gravity - 2.88
Moisture<br> content %<br> (w/w) 5
Comminution Characteristics
Crushing<br> work index (CWi), 75th percentile kWh/t 23.3
JK<br> SMC test parameters (A x b), 25th percentile - 48
Bond<br> ball mill work index (BWi), 75th percentile kWh/t 12.7
Abrasion<br> index (Ai), 75th percentile g 0.048
Crushing
Crushing<br> circuit product size, P80 mm 125
Stockpile<br> live capacity at reclaim rate h 16
Grinding
Grinding<br> circuit product size, P80 μm 75
Gold Leaching
Pre-leach<br> thickener underflow pulp density %<br> solids (w/w) 50
Pre-leach<br> thickener unit area settling rate t/m^2^/h 0.5
Leach<br> residence time h 48
Gold<br> leach extraction % 95.0
Silver<br> lech extraction % 64.0
Leach<br> sodium cyanide (NaCN) addition kg/t<br> leach feed 5.1
Leach<br> lime addition kg/t<br> leach feed 1.3
Counter-current<br> decantation thickener underflow solids concentration %<br> (w/w) 50
CCD<br> wash ratio m³<br> wash water: m³ liquid in CCD feed slurry 3.5
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Description Units Value
--- --- ---
Cyanide Destruction and Tailings Management
Cyanide<br> detoxification residence time min 90
Cyanide<br> detoxification solids concentration %<br> solids (w/w) 40
Circuit<br> feed cyanide concentration, design mg/L<br> CNWAD 530
Circuit<br> feed cyanide concentration, design mg/L<br> CNWAD 460
Circuit<br> discharge cyanide concentration mg/L<br> CNWAD 5
Cyanide<br> detoxification SO2 addition rate, nominal g<br> SO2/g CNWAD 5
Cyanide<br> detoxification SO2 addition rate, design g<br> SO2/g CNWAD 3.5
Tailings<br> thickener unit area thickening rate t/m^2^/h 0.5
Tailings<br> thickener underflow pulp density %<br> solids (w/w) 50

Table 17-5:       MajorEquipment Sizing

Description Specification Size
Crushing
Primary<br> crusher Gyratory<br> crusher <br><br> (Superior MK-III 50-65 or equiv.) 1<br> x 550 kW
Grinding
SAG<br> mill 10.36<br> m Dia. x 6.10 m EGL 1<br> x 14,000 kW
Ball<br> mill (overflow discharge) 7.32<br> m Dia. x 10.06 m EGL 1<br> x 10,000 kW
Ag/Pb Flotation
Rougher<br> flotation cell Tank,<br> forced-air 6<br> x 305 m^3^
Cleaner<br> 1 flotation cell Tank,<br> forced-air 4<br> x 37 m^3^
Cleaner<br> 2 flotation cell Tank,<br> forced-air 2<br> x 37 m^3^
Cleaner<br> 3 flotation cell Tank,<br> forced-air 2<br> x 23 m^3^
Concentrate<br> regrind mill Horizontal<br> High Speed Stirred mill 1<br> x 1,500 kW
Ag/Pb Concentrate Handling
Concentrate<br> thickener High-rate 1<br> x 10 m Dia.
Concentrate<br> filter Pressure<br> filter 1<br> x 109 m^2^
Ag/Zn Flotation
Rougher<br> conditioning cell Tank,<br> forced-air 3<br> x 80 m^3^
Rougher<br> flotation cell Tank,<br> forced-air 4<br> x 305 m^3^
Cleaner<br> 1 flotation cell Tank,<br> forced-air 4<br> x 37 m^3^
Cleaner<br> 2 flotation cell Tank,<br> forced-air 2<br> x 37 m^3^
Cleaner<br> 3 flotation cell Tank,<br> forced-air 2<br> x 23 m^3^
Concentrate<br> regrind mill Horizontal<br> High Speed Stirred mill 1<br> x 1,500 kW
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Description Specification Size
--- --- ---
Ag/Zn Concentrate Handling
Concentrate<br> thickener High-rate 1<br> x 12 m Dia.
Concentrate<br> filter Pressure<br> filter 1<br> x 94 m^2^
Gold Leaching
Pre-leach<br> thickener High-rate 1<br> x 18 m Dia.
Leach<br> tanks - 6<br> x 1,941 m^3^
CCD<br> thickener - 5<br> x 17 m Dia.
Clarifier - 1<br> x 34 m Dia.
Cyanide<br> detoxification tanks - 2<br> x 151 m^3^
17.3.1 Phase 1 Design (Years 1-5)
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17.3.1.1 Primary Crushing
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The primary crushing circuit, located near the ROM blending stockpiles, is designed for 6,570 operating hours per year, corresponding to 75% availability. Run-of-mine (ROM) feed material, with a maximum lump size of 800 mm and P80 of 500 mm, is directly fed into the primary crusher dump pocket by 140 t haul trucks. There will be a rock breaker adjacent to the primary crusher for handling oversized material.

The single-stage open-circuit gyratory crusher (550 kW) discharges into a discharge vault fitted with a reclaim apron feeder. The product from the primary crushing circuit has a design P80 of 125 mm. The apron feeder discharges onto the primary crusher discharge conveyor, which then transfers material to the crushed material stockpile feed conveyor, conveying the crushed material from the primary crushing area to the crushed material stockpile.

For the expansion phase, the primary crushing circuit will be twinned.

Major equipment in this area will include:

· Primary crusher with dump pocket and discharge vault
· Primary crusher discharge apron feeder and discharge conveyor
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· Crushed material stockpile feed conveyor
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17.3.1.2 Stockpile and Reclaim
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The stockpile and reclaim circuit are designed for 7,998 operating hours per year, corresponding to 91.3% availability, at a nominal reclaim rate of 1,000 t/h.

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Crushed material from the primary crusher will be transported by conveyor to a single conical coarse material stockpile. The stockpile has been designed with approximately 16 hours of live capacity at the nominal process plant throughput, providing a buffer between the crushing and grinding circuits and allowing continued mill operation during primary crusher shutdowns.

Material will be withdrawn from the coarse material stockpile via reclaim feeders located within a reclaim tunnel and discharged onto the SAG mill feed conveyor. The SAG mill feed conveyor will subsequently transfer the reclaimed material to the SAG mill feed chute for grinding.

Major equipment in this area will include:

· Reclaim apron feeders (2 duty and 1 standby)
· SAG mill feed conveyor
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17.3.1.3 Grinding
--- ---

The grinding circuit consists of a SAG mill circuit followed by a ball mill circuit in closed configuration with cyclones, designed for 7,998 operating hours per year, corresponding to 91.3% availability, at a nominal throughput of 1,000 t/h.

The circuit is sized based on a grinding circuit feed size (F80) of 125 mm and a circuit product size (P80) of 75 μm. SAG mill slurry will discharge onto a rubber-lined trommel screen with trommel oversize discharging to a pebble recycle conveyor for reprocessing in the SAG mill. SAG mill sizing is based on drop-weight test data supplied by the client from metallurgical test work held within its internal company database. These data are considered suitable for preliminary design purposes; however, they were not generated from samples collected from the Carangas Project. To address the uncertainty associated with the absence of project-specific coarse material competency test work, the design includes an increased SAG mill design factor to provide additional capacity and operating flexibility. This includes the ability to redirect a portion of the cyclone underflow to the SAG mill, if required, to assist with throughput debottlenecking if higher-competency mineralized rock is encountered.

The trommel undersize will combine with the ball mill discharge in the cyclone feed pumpbox where the slurry will be diluted to the desired pulp density with process water and pumped to the cyclone cluster. Flotation reagents lime and zinc sulfate will also be added to the cyclone feed pumpbox. Overflow from the cyclones at 30% solids w/w will report to a trash screen followed by the lead-silver rougher flotation conditioning tank. Cyclone underflow will return to the ball mill directly for further size reduction.

Major equipment in this area includes:

· SAG mill
· Pebble recycle conveyor
· Ball mill
· Primary cyclone cluster
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· Cyclone overflow trash screen
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17.3.1.4 Silver-Lead Flotation
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The silver-lead flotation circuit is designed for 7,998 operating hours per year, corresponding to 91.3% availability, at a nominal throughput of 1,000 t/h. Equipment sizing is based on flotation testwork conducted by ALS Metallurgy (Kamloops).The flotation circuit consists of rougher flotation, followed by rougher concentrate regrinding and three stages of cleaning to produce a saleable silver-lead concentrate grading a minimum of 2000 g/t silver and 24.0% lead.

Cyclone over flow will report to a conditioning tank where lime and collector will be added. Conditioned slurry will then flow by gravity to a bank of conventional forced-air tank cells at a nominal density of 30% w/w where frother will be introduced. The combined rougher concentrate will be collected and pumped to regrinding, while the tailings will be pumped to the lead-silver tailings thickener.

The regrind circuit design consists of a cyclone cluster and a horizontal high-speed stirred regrind mill operating in open circuit. Slurry from the regrind feed pumpbox will be pumped to the cyclones to increase the solids density of the feed to the regrind mill; underflow from the cyclone will report to the regrind mill. Regrind mill discharge and cyclone overflow will report to the lead-silver cleaner 1 conditioning tank where lime and collector will be added. The combined regrind circuit discharge targets a product size P80 of 20 µm.

The lead-silver cleaner circuit design consists of three sequential stages of cleaning, utilizing banks of conventional forced-air tank cells. The first stage will be dosed with frother. The flotation concentrates will flow from the first stage through to the third, and concentrate from the third stage will report to the lead-silver concentrate thickener. Flotation tailings will flow counter-currently to the concentrate, and the first cleaner tailings will report to the lead-silver tailings thickener.

Major equipment in this area includes:

· Lead-silver rougher conditioning tank
· Lead-silver rougher flotation cells
· Lead-silver regrind mill with regrind cyclone cluster
· Lead-silver cleaner conditioning tank
· Lead-silver cleaner flotation cells.
17.3.1.5 Zinc Flotation
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The zinc flotation circuit is designed for 7,998 operating hours per year, corresponding to 91.3% availability, at a nominal throughput of 1,000 t/h. Equipment sizing is based on flotation testwork conducted by ALS Metallurgy (Kamloops).The flotation circuit consists of rougher flotation, followed by rougher concentrate regrinding and three stages of cleaning to produce a saleable zinc concentrate grading a minimum of 45.8% zinc.

Lead-silver tailings thickener underflow is pumped to the zinc rougher conditioning tanks. In first tank, lime and dilution water is added followed by the second tank where copper sulfate is added and finally the third tank where collector is added. The conditioned slurry will then flow to the zinc rougher flotation cell where frother is added. The combined rougher concentrate will be collected and pumped to regrinding, while the tailings will be pumped to the tailings thickener.

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The regrind circuit design consists of a cyclone cluster and a horizontal high-speed stirred regrind mill operating in open circuit. Slurry from the regrind feed pumpbox will be pumped to the cyclones to increase the solids density of the feed to the regrind mill; underflow from the cyclone will report to the regrind mill. Regrind mill discharge and cyclone overflow will report to the zinc conditioning tanks. The combined regrind circuit discharge targets a product size P80 of 20 µm. Copper sulfate will be added to first zinc cleaner conditioning tank and collector will be added to second zinc conditioning tank.

The zinc cleaner circuit design consists of three sequential stages of cleaning, utilizing banks of conventional forced-air tank cells. The first stage will be dosed with frother. The flotation concentrates will flow from the first stage through to the third, and concentrate from the third stage will report to the zinc concentrate thickener. Flotation tailings will flow counter-currently to the concentrate, and the first cleaner tailings will report to the zinc rougher conditioning tank.

Major equipment in this area includes:

· Zinc rougher conditioning tank
· Zinc rougher flotation cells
· Zinc regrind mill with regrind cyclone cluster
· Zinc cleaner conditioning tank
· Zinc cleaner flotation cells
17.3.1.6 Silver-Lead Concentrate Handling
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The silver-lead concentrate handling circuit is designed for 7,998 operating hours per year, corresponding to 91% availability, with concentrate filtration operating 7,796 hours per year, or 89% availability, to accommodate batch filter press operation.

The design of each concentrate dewatering circuit consists of thickening and filtration equipment required to dewater the lead-silver and zinc concentrates prior to loadout and shipment. Each cleaner 3 concentrate stream will report to a dedicated high-rate thickener, where flocculant will be added to assist in the settling of the solids. The lead-silver concentrate thickener overflow will report to process water tank 1 while zinc concentrate thickener overflow will report to process water tank 2. Concentrate thickener underflows will be fed to dedicated filter feed tanks each with a residence time of 24 hours.

The lead-silver thickener underflow will report to a dedicated concentrate filter at a nominal pulp density of 60% w/w. The vertical plate and frame filter pressed will discharge filter cake at onto the lead-silver concentrate stockpile.

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The zinc thickener underflow will report to a dedicated concentrate filter at a nominal pulp density of 60% w/w. The vertical plate and frame filter pressed will discharge filter cake at onto the zinc concentrate stockpile.

The lead-silver and zinc concentrate handling circuits will be physically separated from one another to prevent cross contamination. The concentrates will be reclaimed from the stockpiles by a front-end loader and is then loaded into containerized highway haulage trucks for off-site transport.

Major equipment in this area includes:

· Lead-silver concentrate thickener
· Lead-silver concentrate filter feed tank and agitator
· Lead-silver concentrate filter
· Zinc concentrate thickener
· Zinc concentrate filter feed tank and agitator
· Zinc concentrate filter
17.3.2 Phase 2 Design (Years 6-8)
--- ---

Equipment in Phase 1 will be duplicated, resulting in two dedicated 8.0 Mt/a concentrators resulting in a total mill throughput of 16.0 Mt/a.

17.3.3 Phase 3 Design (Years 9-16)

The crushing, stockpile and reclaim, grinding and rougher flotation circuit with regrind mill from one of the 8.0 Mt/a silver-lead-zinc concentrators is used to process the gold mineralised material.

17.3.3.1 Gold Leaching

The gold leaching circuit is designed for 7,998 operating hours per year, corresponding to 91.3% availability, at a nominal throughput of 1,000 t/h. Equipment sizing is based on Base Metallurgical Laboratories Ltd. leaching testwork.

The gold leaching circuit consists of a pre-leach thickener, a leach pre-aeration stage, and six cyanide leach tanks. The pre-leach thickener is used to recover cyanide-free water ahead of the leaching circuit, with thickener overflow reporting to the cyanide-free process water tank. Thickener underflow, at 50% solids by weight, is pumped to the leach pre-aeration tank, where gold barren cyanide solution from the Merrill-Crowe circuit is used to adjust solids concentration. Lime and oxygen are added to the pre-aeration tank to adjust slurry chemistry and increase dissolved oxygen levels prior to cyanidation. Following pre-aeration, sodium cyanide is added and leaching is conducted in six mechanically agitated and aerated tanks arranged in series.

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Major equipment in this area include:

· Pre-leach thickener
· Pre-aeration tank with agitator
· Six leach tanks with agitators
17.3.4 Counter-Current Decantation
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The gold recovery and carbon regeneration circuit is designed for 8,059 operating hours per year, corresponding to 91.3% availability.

Leached slurry will report to the first of five counter-current decantation (CCD) thickeners. Thickened solids will be pumped sequentially through the CCD train, counter-current to the wash solution flow. Barren solution from the Merrill-Crowe circuit will be used to wash the leached solids and will be introduced with the slurry feed to the fifth CCD thickener. Overflow solution will flow by gravity counter-currently through the CCD train, with overflow from the first thickener reporting to the pregnant solution clarifier. The CCD and clarifier circuit is designed to limit the residual dissolved silver and gold concentrations in the fifth thickener underflow solution to a maximum of 0.5 mg/L and 0.05 mg/L, respectively. Clarifier overflow will report to the pregnant solution tank for subsequent feed to the Merrill-Crowe circuit, while clarifier underflow will be pumped, as required, back to the bulk leaching circuit. Flocculant will be added, as required, to enhance solids settling and improve overflow clarity.

Major equipment in this area include:

· Five counter-current decantation thickeners
· One pregnant solution clarifier
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17.3.5 Merrill Crowe Circuit
--- ---

The gold recovery and carbon regeneration circuit is designed for 8,059 operating hours per year, corresponding to 91.3% availability.

Pregnant solution overflow from the clarifier will report to the pregnant solution tank prior to precious metal recovery by zinc cementation in the Merrill-Crowe circuit. Solution from the pregnant solution tank will be pumped through polishing filters to further reduce total suspended solids, followed by treatment in a deaeration tower to reduce dissolved oxygen content. Deaerated solution from the tower will then be contacted with zinc powder to precipitate dissolved gold and silver. The resulting precipitate-bearing slurry will be pumped to two plate-and-frame precipitation filters for recovery of the precious-metal-rich precipitate.

Filter solids will be discharged and transferred manually within the refinery for subsequent doré production. Filtrate from the precipitation filters will report to the barren solution tank. Barren solution will be distributed to the leaching circuit and CCD circuit for use as wash solution, with a bleed stream directed to the detoxification circuit, as required.

Antiscalant will be added to the pregnant and barren solution tanks to reduce the potential for scale formation within the Merrill-Crowe circuit. Diatomaceous earth will be added to the clarifying and precipitation filters to improve filtration performance. Zinc powder will be added to the precipitation filter feed to achieve greater than 99% precipitation of gold and silver. Lead nitrate will be added to the zinc cone to improve zinc efficiency and reduce the potential for zinc passivation, which could inhibit the cementation reaction. Sodium cyanide will also be added to the zinc cone to support effective gold and silver precipitation.

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Major equipment in this area will include:

· one pregnant solution tank
· two rotating disc clarifying filters
· one deaeration tower
· two plate-and-frame precipitation filters
· one barren solution tank.
17.3.6 Refinery
--- ---

Zinc precipitate recovered from the Merrill-Crowe circuit will be dried prior to batch smelting to produce silver-gold doré. The dried precipitate will be blended with fluxes and charged to an electric melting furnace. During smelting, the fluxes will react with base-metal constituents and other impurities to form oxides, which will report to the slag phase and separate from the molten precious-metal phase.

The molten metal will be poured into moulds to produce doré bars. Doré bars will be cleaned, sampled and assayed, stamped, and stored in a secure vault prior to periodic shipment to market.

The gold room will be provided with suitable ventilation and off-gas handling systems to maintain a safe working environment and to control furnace fumes and dust generated during smelting and materials handling. The ventilation system will include hoods, enclosures, ducting, and fans designed to meet applicable local regulations and guidelines. Mercury present in the precipitate, if any, may volatilize during smelting and will be captured in the mercury abatement system, which will form part of the gold room ventilation system and will include sulfur-impregnated carbon.

A gold room sump, equipped with a precious-metals trap, will be installed to collect hose-down water and spillage. Collected material will be returned to the clarifier feed box.

The refinery will be supplied as a vendor package and will include the following major equipment:

· flux dosing and mixing system
· electric melting furnace
· doré scale and secure storage vault
· slag handling equipment
· dust collection and mercury abatement system.
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17.3.7 Cyanide Detoxification and Tailings Management
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The cyanide destruction and tailings management circuit is designed for 8,760 operating hours per year, corresponding to 91.3% availability.

The circuit consists of two parallel cyanide detoxification tanks, followed by a tailings thickener. The dewatered tailings slurry reports to the tailings storage facility (TSF), while recovered process water is recycled to the cyanide containing process water tank for reuse.

The tailings from the CCD thickener 5 are fed to the two parallel cyanide detoxification tanks, where the slurry remains for a total of 90 minutes. The circuit is designed to decrease weak acid dissociable cyanide (CNWAD) concentration from 560 mg/L to no more than 5.0 mg/L. Cyanide destruction is accomplished using the sulfur dioxide/air method, with detoxification reaction carried out at a pH of 8–9, maintained by the addition of lime. Sulfur dioxide (SO2) is supplied in the form of sodium metabisulfite (SMBS), while copper sulfate in solution serves as a catalyst. The cyanide destruction tanks are equipped with oxygen sparging and mechanical agitation to ensure thorough mixing. Following detoxification, the slurry is pumped to the tailings thickener.

Tailings thickening is performed using a single a high-rate thickener. The slurry is thickened to recover process water and reduce tailings volume prior to discharging to the TSF. Flocculant is added to the thickener feed stream to enhance settling. The thickener overflow reports to the cyanide containing process water tank, while the underflow is collected at 50% solids and pumped to the TSF.

17.4 Energy, Water and Process Materials Requirements
17.4.1 Process Materials
--- ---

The estimated annual consumption based on nominal usage for major plant reagents is summarised in Table 17-6 and Table 17-7.

Table 17-6: ReagentsConsumption for Concentrate Plant
Reagent Phase 1 consumption (t/a) for the 8.0 Mt/a concentrator Phase 2 consumption (t/a) for the 16.0 Mt/a concentrator
--- --- ---
Lime 9,120 18,240
Collector AP3418A 208 416
Collector Aero 404 32 64
Collector SIPX 520 1,040
MIBC 240 480
Frother W31 80 160
Zinc sulfate 10,400 20,800
Copper sulfate 1,200 2,400
Antiscalant 0.08 0.16
Flocculant 425 850
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Table 17-7: Reagents Consumption for Gold Plant
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Reagent Annual Consumption (t/a)
--- ---
SIPX 144
MIBC 400
Lime 4,186
Sodium cyanide - concentrate leach 1,134
Lead nitrate 144
Diatomaceous earth 839
Zinc powder - gold and silver precipitation 68
Flux additives - Borax 100
Flux additives - Nitre 10
Flux additives - Sodium carbonate 50
Flux additives - Silica 50
Antiscalant 64
SMBS 3,087
Copper sulfate 51
Oxygen 3,673
Flocculant 441
17.4.2 Water Requirements
--- ---
17.4.2.1 Raw Water
--- ---

Raw water will be provided to the plant and stored in a raw water storage tank where it will be distributed to the various users across the plant site such as reagent preparation, gland seal water, potable water, and plant make-up. Long term consumption requirements will be approximately 3,400,000 m^3^ per annum per concentrator and 4,300,000 m^3^ per annum for the gold plant.

17.4.2.2 Process Water

Process water for the silver-lead and zinc concentrators will be provided in two separate water tanks to ensure that dissolved reagents do not interfere with the respective silver-lead and zinc flotation processes. Process water prior to zinc flotation (lead-silver flotation tailings thickener and lead-silver concentrate thickener overflow) will report to process water Tank 1, with remaining process water reporting to process Tank 2.

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Process water for the gold plant will be provided in two separate water tanks to ensure that dissolved sodium cyanide does not interfere with the gold flotation processes. Process water prior to gold flotation (gold rougher tailings thickener and pre-leach thickener overflow) will report to process water Tank 1, with remaining process water reporting to process Tank 2.

Raw water will be provided as required for makeup. Process water will be stored in a storage tank before being pumped across the plant site to the various end users.

17.4.3 Air Requirements

High-pressure air at 750 kPa is produced by compressors to meet plant requirements. The high-pressure air supply is dried and used to satisfy both plant air and instrument air demand. Dried air is distributed via the air receivers located throughout the plant. Compressed air filtration demand is generated by dedicated filter compressors.

17.4.4 Oxygen Requirements

Oxygen gas for the cyanide leaching and detoxification circuits in Phase 3 for the gold plant will be generated from an on-site oxygen plant. The oxygen plant will employ vacuum swing adsorption (VSA) technology and will produce an oxygen stream of 92% purity at 6 bar (g).

17.4.5 Power requirements

The estimated installed load for the crushing plant and process plant is 41.0 MW during the Phase 1, 82.0 MW during the Phase 2 and 79.4 MW for Phase 3 (38.4 MW for the gold plant). The estimated nominal operating demand is 28.6 MW during the Phase 1, 57.2 MW during for Phase 2 and 55.4 MW for Phase 3 (26.8 MW for the gold plant).

17.4.6 Personnel Requirements

The estimated personnel requirements for the process operations are 102 people for Phase 1 and 189 people for Phase 2 and 3. General and administrative personnel requirements will be 57 people for Phase 1 and 92 people for Phase 2 and 3.

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18 Project Infrastructure
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18.1 Introduction
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The Carangas Project will make use of current existing off-site infrastructure such as roads, transmission lines, and port facilities where applicable. Additional on-site infrastructure will be required to be built during the construction phase of the Project.

The on-site infrastructure for the Project includes:

· Access road
· Site roads and laydowns
--- ---
· WRSF
--- ---
· Oxide stockpile
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· Low-grade stockpile
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· Gold zone stockpile
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· ROM stockpile
--- ---
· Operations camp
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· Auxiliary buildings
--- ---
· Mine office and mine dry
--- ---
· Truck shop, maintenance shop, and warehouse
--- ---
· Mine rescue facilities
--- ---
· Fuel storage and distribution facilities
--- ---
· Power plant and electrical distribution
--- ---
· TSF and water management structures
--- ---
· Explosives magazine.
--- ---

The off-site infrastructure required for the project includes:

· National Highway 12 from the City of Oruro to the Town of Sabaya
· Site access road from Sabaya to site
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· 230 kV Overhead transmission line from Pagador to Carangas.
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Figure 18-1: General Arrangement
--- ---

Source: Ausenco, 2026

18.2 Site Access

The Project site is currently accessible by vehicle by driving approximately 200 km via National Highway 12 from the City of Oruro to the Town of Sabaya. From Sabaya, a 6 km paved road followed by a 30 km gravel road leads to the Carangas site.

The concentrate products will be loaded on bulk haulage highway trucks and transported via the 36 km access road to Sabaya. The bulk haulers will then follow Bolivian Ruta Nacional (RN) 12 for 38 km to the Bolivian-Chilean border at the Colchane border crossing. The route will then continue for 233 km via Chilean RN 15 to Huara, Chilean RN 5 to Humberstone, then Chilean RN 16 to the port of Iquique. The total distance from site to port is 307 km. The bulk concentrate will then be unloaded, stockpiled, assayed, and loaded on ships.

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The port facilities are considered a public port, and are owned and operated by a Empresa Portuaria Iquique, a Chilean state-owned entity.

18.3 Built Infrastructure
18.3.1 On-site Roads
--- ---

The Project requires multiple haul roads to be constructed. The haul roads will connect the Carangas open pit with the crusher, WRSF, Oxide Stockpile (OXSP), Low-Grade Stockpile (LGSP), truck shop, and TSF. On-site access roads will also need to be constructed. The Carangas Access Road, which connects the site to the Town of Sabaya will require upgrades. Access to the explosives magazine will be located in the northwest area of the claims block.

18.3.2 Process Plant Building

The process plant buildings will be pre-engineered buildings and will house the grinding, flotation, reagent, concentrate filtration, gold plant, and load-out areas. Additional support buildings will be located on or adjacent to the process plant site.

18.3.3 Accommodation

A construction camp capable of accommodating 590 people will be assembled from prefabricated modules during the construction phase. Additional modules will be added if required. The camp will include a central complex with dining, kitchen, and recreational facilities. For operations, the construction camp will be updated to accommodate 295 operations staff during the initial production phase and expanded to accommodate 490 people during the first expansion phase.

18.3.4 Support Buildings

Additional support buildings will be required. These include the explosives magazine, primary crushing, reclaim tunnels, gatehouse, e-rooms, plant offices, mine offices, change houses, truck shop, warehouse/workshop, fuel storage and distribution, core storage and yard, potable and sewage treatment plants, medical and emergency, and laboratory buildings.

18.4 Stockpiles

There are two stockpiles located on the site; the OXSP and LGSP. The OXSP include the storage of all oxide material, to be blended with the sulfide material and fed into the mill. The LGSP includes the storage of all gold zone material mined prior to the start of the gold plant. This material is fed into the gold plant once the plant is up and running.

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18.5 Tailings Storage Facility (TSF)
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The primary objectives of the TSF are the secure confinement of slurry tailings and the protection of the regional groundwater and surface water during operations and post-closure. The following have been considered in the design of the TSF and water management:

· Geosynthetic lining of the embankment’s upstream slope to limit seepage (the requirement for<br>a basin liner will be reviewed during future design and environmental assessment phases).
· Control, collection, and removal of water from the facility during operations for recycling as process<br>water to the maximum practical extent.
--- ---

A total of 251.0 Mt of tailings will be stored in the TSF. Construction of the TSF has been divided into multiple phases: Phase 1, the starter facility, will store 16.1 Mt of tailings, and the ultimate facility will store an additional 235,025.2 Mt of tailings, all of which will be pumped by pipeline from the process plant to the TSF. The TSF is located at the northwest side of a large valley 2.6 km northwest of the process plant. The final TSF embankment and impoundment basins will occupy an ultimate footprint of 3.77 Mm^2^ (337 ha).

The general arrangement of the TSF is shown in Figure 18-2.

Figure 18-2: TSF General Arrangement

Source: Ausenco, 2026

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The primary design objectives for the TSF are the secure confinement of tailings, the subaerial deposition of tailings, and the protection of regional groundwater and surface water during mine operations and in the long-term post-closure period.

The design of the TSF and associated water management facilities has taken into account the following:

· Staged development of the facility over the life of the project
· Flexibility to accommodate operational variability in the tailings (plant shutdowns, deposition variability,<br>and placement during variable climate conditions)
--- ---
· Control, collection, and removal of contact water from the facility during operations for reuse as process<br>water to the maximum practical extent.
--- ---

The design criteria for the TSF consider the following requirements for slurry tailings:

· Tailings storage requirement: approximately 251.0 Mt
· Tailings dry density: 1.45 t/m^3^
--- ---
· The tailings embankment will be built with NAG waste rock with an upstream geomembrane liner, low permeability<br>soil layer, and a filter layer between the low permeability and filter layers
--- ---
· Subaerial deposition of tailings
--- ---
· Limiting watershed disturbance
--- ---
· Limiting impacts on wildlife and fisheries resources.
--- ---

The starter TSF has been designed to contain the first two years of slurry tailings production of 16.1 Mt (Figure 18-3) and the ultimate TSF has been designed to contain the LOM tailings production of 304.3 Mt (Figure 18-4).

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Figure 18-3: TSF Starter Facility
--- ---

Source: Ausenco, 2026

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Figure 18-4: TSF Ultimate Facility
--- ---

Source: Ausenco, 2026

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18.5.1 TSF Design and Construction
--- ---

The TSF footprint will be cleared for foundation preparation and embankment construction. Basin preparation will include the removal of topsoil and soft overburden beneath the embankment foundation prior to placing waste rock. A foundation drainage network will be developed at the base of the embankment, with drainage gravel and a dual-wall HDPE pipe wrapped in nonwoven geotextile fabric.

The starter embankment will be constructed using NAG waste rock, with an upstream low-permeability layer and a filter layer above the waste rock embankment, and a geomembrane liner anchored several meters into the alluvium. NAG waste rock and till will be transported to the TSF by haul trucks and placed in a borrow-source stockpile for embankment construction. The waste rock will be spread and compacted with dozers and compactors into 1 m lifts. The embankment will be constructed with overall 2.5:1 (H:V) downstream slope, an upstream slope of 2.0:1 (H:V) and a 15 meter-wide crest, which provides a stable configuration under both static and dynamic loading based on the stability analyses (Figure 18-5). Subsequent raises will utilize the downstream construction method.

Figure 18-5: Cross-section for CDSF Embankment

Source: Ausenco, 2026

The design standards for the TSF is based on the relevant federal and international guidelines for the construction of mining tailings storage facilities in Bolivia. Dam breach assessment and dam hazard classification were excluded from the scope of work of this study. However, regulations and guidelines, such as the Inflow Design Flood (IDF) and the Earthquake Design Ground Motion (EDGM), and the Bulletin - Application of Dam Safety Guidelines to Mining Dams (CDA, 2019), were used to speculate on the dam hazard classification and suggested minimum target levels for some design criteria. Based on the expected downstream inundation area of the TSF, the consequence of a dam failure is “Very High” for the TSF, according to CDA 2007 (revised 2013). Therefore, the facility was designed following the recommended parameters in these guidelines.

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To be conservative, the inflow design flood (IDF) used for the TSF design during operations and post-closure is the probable maximum flood (PMF) for a “Very High” dam classification. The TSF spillway is designed to accommodate the PMF. The perimeter diversion channels are designed to capture runoff above the channels for the 100-year, 24-hour storm event and safely redirect it around the TSF. EDGM parameters for the TSF have been determined using the Bolivian seismic hazard map, which covers only a 1:475 return period. The design earthquake for the TSF is the maximum credible earthquake (MCE) for a “Very High” dam classification post-closure. Therefore, the 1:475 peak ground acceleration (PGA) was multiplied to obtain the MCE, based on a similar project with the full spectrum of seismic return periods.

18.5.2 Stability Analysis

A section through the highest portion of the embankment was selected as the critical section. Stability of the embankment was assessed using the limit-equilibrium modelling software SLIDE2v9.03 (Rockscience, 2021). Analyses were undertaken for both static and pseudo-static (earthquake loading) conditions, with the calculated factors of safety (FoS) exceeding the minimum required values in accordance with CDA guidelines: 1.5 FoS for static and 1.0 FoS for pseudo-static. The TSF embankment is designed to withstand potential dynamic displacement without release of tailings during the maximum design earthquake event. The embankment stability analyses exceeded both static and pseudo-static CDA guidelines.

18.5.3 Geotechnical Instrumentation and Monitoring

Instrumentation and monitoring will be required to assess the performance of the embankment and must be incorporated in the next phase of the study. Vibrating-wire piezometers will be installed to monitor pore pressure in the TSF fill materials, and slope inclinometers and survey monuments will be installed in the embankment to monitor slope movement and deformation.

18.5.4 TSF Closure

The ultimate TSF will occupy approximately 337 ha at the northwest limit of the watershed. The spillway will be constructed in bedrock. The embankment will be covered with soil-vegetative cover to protect against erosion. A waste rock cover and soil-vegetative cover will be placed over the tailings to reduce infiltration and promote surface runoff to the spillway.

18.6 Waste Rock Storage Facility

The waste rock management plan includes the storage of all the mined material below the NSR cut-off value of $13.50 (waste). The WRSF is designed to store 87.92 million cubic meters of material. The slope of the ultimate facility is designed at 3H:1V with a maximum height of 90 m and a maximum elevation of 4,010 masl. The ultimate WRSF design is illustrated in Figure 18-6.

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Figure 18-6: Ultimate WRSF Design
--- ---

Source: Ausenco, 2026

18.7 Power and Electrical

The Project requires a dedicated electrical power supply to support the process plant, on-site infrastructure, and select mining equipment. The power system is designed to deliver up to 82 MW of peak electrical load over the LOM. The average annual operating load for the process plant and on-site infrastructure is estimated at 57 MW. The average annual operating load for mining operations is estimated at 6 MW.

Power will be supplied by a new 230 kV transmission line connecting the Project site with Pagador substation.

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The power cost used for this study is US$0.06/kWh and was provided to NPM by Empresa Nacional de Electricidad (ENDE), the major national electricity company in Bolivia.

18.8 Fuel

On-site diesel storage is designed to accommodate two weeks of fuel. Weekly consumption for the mining equipment is estimated at 385,000 L. On-site storage capacity is 850,000 L for the initial phase and 1,700,000 L for the expansion phase. These capacities include considerations for light vehicles and mobile equipment. Diesel fuel will be delivered by truck via National Highway 12.

The fuel price used for this study is US$1.41 per liter of diesel. This is based on the current diesel cost for the exploration activities and was provided by NPM.

18.9 Water Supply and Management
18.9.1 Water Supply
--- ---

Water is assumed to be sourced through drilling in the project vicinity upstream of the pit. This will serve to ensure the pit has minimal inflows and provide water for the processing facility. Two small local streams run through the property with a flow rate of approximately 20 L/s for each of them in the dry season. This water supply was adequate for the project's water consumption during the exploration and drilling stage. The Project expects to source additional water sources through drilling to identify new and utilizing existing wells. The adequacy of the water supply as planned is to be supported by hydrology studies that are planned to be completed.

The backup source for water supply has been identified; however, further studies are required to confirm its availability and accessibility.

At the site, fresh water will be stored in two carbon steel tanks, each 5 m diameter x 5 m height, close to the process plant, where water will be filtered and treated for process water or potable water use. Additionally, a potable water tank of the same size will be required.

Depending on its service, water will be distributed from the process plant to all required areas. Firewater will be stored in the freshwater tank, which will have a dedicated volume and two hours of storage capacity.

18.9.2 Water Management
18.9.2.1 Water Management for Open Pit, Stockpiles, Camp/Admin, and Process Areas
--- ---
18.9.2.1.1 Objectives
--- ---

The objective is to effectively separate contact water (from disturbed mining areas) and non-contact water, direct flows to collection ponds, and provide adequate hydraulic capacity for the 100-year, 24-hour design storm. The other objectives are:

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· Calculate design peak flows for channels and structures (using HEC-HMS).
--- ---
· Size open channels, culverts, and contact water ponds based on the site layout.
--- ---
· Ensure compliance with environmental requirements through proper contact vs. non-contact water segregation.
--- ---
· In Year 1, the Carangas River will be moved to the east to accommodate the expansion of the open pit over<br>the LOM.
--- ---
18.9.2.1.2 Modelling Approach
--- ---

The design storm adopted for all hydraulic structures is the 100-year, 24-hour precipitation event of 45 mm. Monthly precipitation and evaporation data from the Santos Station and Molina Carpio records were used for the water balance component.

Peak discharges for the channels were developed using the HEC-HMS hydrologic model. The model incorporated SCS Curve Number loss method, Clark Unit Hydrograph transformation, and basin-specific parameters including time of concentration and storage coefficient derived from longest flow path analysis and watershed characteristics derived from AutoCAD drawing [U.S. Army Corps of Engineers, Hydrologic Engineering Center (USACE HEC), 2023]. Runoff coefficients of 0.8 for the open pit and stockpiles and 0.9 for the camp and administration areas were applied (e.g., Bedient et al., 2019; Ontario Ministry of Transportation, 1997). The resulting peak flows formed the basis for channel sizing.

18.9.2.1.3 Channel and Pond Sizing

Five channels were planned and designed to convey runoff from the respective catchments and mine areas, with a combined total length of 9,781 m (Table 18-1). These include three non-contact channels (MG/LG Non-contact 1 and 2, and Open Pit Non-contact) that discharge to the environment, and two contact channels (Open Pit Contact and MG/LG Contact) that convey water to the respective collection ponds.

All channels were designed as trapezoidal sections with 1.5H:1V side slopes, 0.30 m freeboard, and Manning’s roughness coefficient of 0.04 for grouted riprap. Normal depth, velocity, and Froude numbers were calculated for each channel (Sturm, 2010). Excavation volumes and lining areas were computed based on the calculated normal depth and freeboard. Contact channels were lined with grouted riprap based on the calculated D50 values and a liner, while non-contact channels use geotextile.

Table 18-1: Channel Design Parameters
Channel Name Q Design <br><br>(m³/s) Length (m) Geometry Slope <br><br>(H:1V) Base B (m) Depth (m) Velocity <br><br>(m/s)
--- --- --- --- --- --- --- ---
MG/LG- Non-contact Water Channel 1 7.1 1,819 Trapezoidal 1.50 2.5 0.76 1.71
MG/LG- Non-contact Water Channel 2 7.7 2,708 Trapezoidal 1.50 2 0.85 1.93
Open Pit- Non-contact Water Chennel 6.4 952 Trapezoidal 1.50 1.5 1.00 1.59
Open Pit-Contact Water Chennel 1.9 1,031 Trapezoidal 1.50 1.5 0.49 0.87
MG/LG-Contact Water Channel 3.2 3,271 Trapezoidal 1.50 1.5 0.68 1.11
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Ponds were sized using the 100-year 24-hour storm runoff volume with a safety factor of 1.25 (Table 18-2).

Table 18-2: Pond Sizing Parameters
Pond Name Contributing<br><br>Area (ha) Runoff <br><br>Volume (m^3^) Safety <br><br>Factor Required <br><br>Storage (m^3^) Avg Depth <br><br>(m)-assumed Est. Surface <br><br>Area (m²)
--- --- --- --- --- --- ---
Pond 1 - Open Pit Contact 49 17,497 1.25 21,871 6 3,645
Pond 2 - Oxide + LG/MG Contact 115 41,225 1.25 51,531 6 8,588
Pond 3 - Camp + Admin 4 1,438 1.25 1,797 3 599
Pond 4 - Process, Truck Shop + MIA Runoff 9 3,245 1.25 4,056 4 1,014

Two 900 mm diameter culverts are included as shown in the layout drawing (Figure 18-2): one on the MG/LG Contact Channel to Pond 2 and one serving the process plant area to Pond 4. Each has a length of 40 m.

18.9.2.1.4 Monthly Water Balance

A monthly water balance was developed using the rational method for runoff, site-specific precipitation, and evaporation (Table 18-3).

Precipitation varies from 0.1 mm in dry months to 89.80 mm in wet months, with high evaporation throughout the year (annual total of 1,313.9 mm). The monthly water balance shows the following annual performance:

· Pond 1 (Open Pit): Annual inflow 101,250 m^3^, annual outflow 96,900 m^3^
· Pond 2 (Oxide + LG/MG): Annual inflow 238,554 m^3^, annual outflow 228,304 m^3^
--- ---
· Pond 3 (Camp + Admin): Annual inflow 8,321 m^3^, annual outflow 7,666 m^3^
--- ---
· Pond 4 (Process Area): Annual inflow 21,126 m^3^, annual outflow 19,950 m^3^
--- ---

The ponds maintain adequate storage with assumed pumping or reuse outflows, demonstrating a net positive water management balance under average conditions (Table 18-4).

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Table 18-3: Monthly Water Balance: Precipitation, Evaporation and PondInflows
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Month Precipitation <br><br>(mm) Evaporation <br><br>(mm) Total Pond 1 <br><br>inflow (m^3^) Total Pond 2 <br><br>inflow (m^3^) Total Pond 3 inflow <br><br>(m^3^) Total Pond 4 <br><br>inflow (m^3^)
--- --- --- --- --- --- ---
Jan 89.8 131.9 34,917 82,266 2,869 7,285
Feb 76.8 113.5 29,862 70,357 2,454 6,231
Mar 44.4 118.4 17,264 40,675 1,419 3,602
Apr 5 98.8 1,944 4,581 160 406
May 0.1 82.9 39 92 3 8
Jun 0.5 72.4 194 458 16 41
Jul 1.6 74.8 622 1,466 51 130
Aug 0.6 93.1 233 550 19 49
Sep 1.3 106.7 505 1,191 42 105
Oct 3.5 128.2 1,361 3,206 112 284
Nov 6.6 147.8 2,566 6,046 211 535
Dec 30.2 145.4 11,743 27,666 965 2,450
Annual 260.4 1313.9 101,250 238,554 8,321 21,126
Table 18-4: Monthly Water Balance: Pond Outflows
--- ---
Month Total Pond 1 outflow (m^3^) Total Pond 2 outflow (m^3^) Total Pond 3 outflow (m^3^) Total Pond 4 outflow (m^3^)
--- --- --- --- ---
Jan 34,436 81,133 2,790 7,151
Feb 29,448 69,382 2,386 6,115
Mar 16,832 39,658 1,348 3,482
Apr 1,584 3,732 101 305
May 0 0 0 0
Jun 0 0 0 0
Jul 349 823 6 54
Aug 0 0 0 0
Sep 117 275 0 0
Oct 894 2,105 35 154
Nov 2,027 4,777 122 386
Dec 11,213 26,418 878 2,303
Annual 96,900 228,304 7,666 19,950

The surface water management system effectively collects contact water from the open pit and stockpiles into dedicated ponds while routing non-contact water safely to the environment. The HEC-HMS modelling provided reliable peak flows, and the monthly water balance confirms that the ponds have sufficient capacity under the design conditions, with outflows maintaining storage levels throughout the year.

For the next study phases, it is recommended to undertake stochastic water balance modelling to evaluate climate variability. Integration with the overall site water balance should also be progressed.

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18.10 Hazard Considerations
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The Carangas area represents a low-to-moderate natural hazard environment. Key considerations for project design and operations include seismic design, surface water management, and cold-weather operational planning. No extreme weather or high-risk geohazards are anticipated to materially constrain project development when appropriate mitigation measures are applied.

18.10.1 Seismic Risk

The Carangas area lies within a region of moderate seismicity, associated with the subduction of the Nazca Plate beneath the South American Plate. While the Bolivian Altiplano is not among the highest seismic-hazard zones in the Andes, seismic loading should be considered in the design of critical infrastructure.

Mitigations

· Apply regionally appropriate seismic design criteria for:
o Tailings storage facilities
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o Dams
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o Process plants
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o Pipelines and power infrastructure
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· Implement site-specific seismic hazard assessments at advanced study stages
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· Incorporate site-specific seismic parameters at advanced study stages.
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18.10.2 Geohazards (Landslides, Rockfall, Debris Flow)
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The project area is characterized by gently undulating Altiplano terrain with limited steep slopes. As a result, regional landslide, rockfall, and debris-flow hazards are generally low. Potential geohazards are primarily associated with engineered slopes and earthworks.

Mitigations

· Limit infrastructure placement in steep or incised local terrain where practicable
· Conduct geotechnical investigations for:
--- ---
o Road cuts
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o Embankments
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o Borrow areas
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· Apply conservative slope design and appropriate drainage controls
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· Monitor engineered slopes during construction and early operations.
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18.10.3 Cryospheric Hazards (Permafrost, Snow, Ice, Freeze–Thaw)
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Permafrost is not expected to be present in the Carangas area. However, cold temperatures and diurnal freeze–thaw cycles are common due to high elevation, and intermittent snowfall may occur.

Mitigations

· Design earthworks and pavements to accommodate freeze–thaw cycling
· Protect exposed pipelines and water infrastructure against freezing
--- ---
· Schedule sensitive construction activities (e.g., concrete placement) during favorable temperature windows
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· Implement basic snow and ice management plans for site access roads.
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18.10.4 Flooding and River-Related Hazards
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The region experiences low annual precipitation, concentrated during a short wet season (generally December to March). Flood hazards are typically localized and associated with short-duration storm events, surface runoff, and ephemeral drainage channels.

Mitigations

· Avoid siting infrastructure within natural drainage channels and low-lying areas prone to ponding
· Design surface water management systems to accommodate seasonal runoff
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· Incorporate flood-freeboard and erosion controls into infrastructure design
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· Maintain drainage features to ensure continued capacity during wet-season events.
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18.10.5 Weather-Driven Hazards (Winter Storms, Heavy Snow, Extreme Precipitation)
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The Carangas area is not subject to cyclones or hurricanes. Weather-related hazards are limited to localized convective storms, occasional snowfall, and cold temperatures that may temporarily affect access or construction efficiency.

Mitigations

· Plan construction schedules to prioritize dry-season activities.
· Design access roads to remain serviceable during wet-season storm events.
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· Establish operational procedures for temporary weather-related interruptions.
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· Maintain weather monitoring to support short-term planning and logistics.
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18.10.6 Avalanche Hazards
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Avalanche hazards are considered negligible due to the absence of steep alpine terrain, sustained snowpack, and high-relief valley systems in the project area.

Mitigations

· No specific avalanche mitigation measures are required.
· Confirm absence of avalanche-prone terrain during site reconnaissance and detailed design.
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18.10.7 Climate-Driven Change
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Long-term climate change in the region may alter precipitation patterns and temperature variability, potentially affecting runoff intensity, water availability, and freeze–thaw conditions over the life of the project.

Mitigations

· Apply conservative assumptions for precipitation and runoff in water management design.
· Periodically review climate data during operations to identify emerging trends.
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· Design infrastructure with sufficient adaptability to accommodate future climate variability.
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· Incorporate climate-resilience considerations in long-term closure planning.
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19 Market Studies and Contracts
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19.1 Market Studies
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No market studies or product valuations were completed as part of this study. Market price assumptions were based on a review of public information, industry consensus, standard practice, and specific information from comparable operations.

Silver-lead and zinc concentrates are widely traded and can be marketed directly from producer to smelter or via third-party concentrate trading entities. It is assumed that the concentrate contains negligible deleterious elements that would impact marketability.

The market for gold doré is widely traded and can be marketed domestically or internationally with significant optionality regarding the final customer. It is assumed that the doré contains negligible deleterious elements that would impact marketability.

A marketing study was not conducted to determine indicative treatment and refining terms. Marketing, refining, and transportation costs, along with payability terms, were informed by preliminary discussions with a major commodity trading company and information from comparable recent studies. The assumed payability terms and off-site costs for each metal are presented in Table 19-1.

Table 19-1: Off-take Term Assumptions
Term Unit Value
--- --- ---
Silver Payability – Ag/Pb Concentrate % 96.5
Lead Payablility – Ag/Pb Concentrate % 95.0
Silver Payability – Zn Concentrate % 70.0
Zinc Payablility – Zn Concentrate % 85.0
Gold Payability – Doré % 99.8
Silver Payability – Doré % 95.0
Transport Cost – Ag/Pb Concentrate US$/wmt 120.00
Transport Cost – Zn Concentrate US$/wmt 120.00
Lead Treatment Cost – Ag/Pb Concentrate US$/dmt 100.00
Zinc Treatment Cost – Zn Concentrate US$/dmt 175.00
Silver Refining Cost – Ag/Pb Concentrate US$/oz 0.50
Silver Refining Cost – Zn Concentrate US$/oz 0.50
Gold Refining, Transport, and Marketing Cost – Doré US$/oz 6.50
Silver Refining, Transport, and Marketing Cost – Doré US$/oz 1.00
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The QP for this section has reviewed the information in this section and, in the QP’s opinion, the data supports the assumptions in this technical report.

19.2 Commodity Price Projections

Project economics were estimated based on long-term flat metal prices of US$45.00/oz Ag, US$3,400/oz Au, US$0.90/lb Pb and US$1.20/lb Zn, which are based on consensus forecasts from various financial institutions.

The QP notes that the pricing used in the cash flow analysis is reasonably aligned with various long-term forward-looking estimates from major international banks.

19.3 Contracts

Currently, there are no contracts for transportation or off-take of any metal products in place, but when they are negotiated, they are expected to be within industry norms. Similarly, there are no contracts currently in place for the supply of reagents, utilities, or other bulk commodities required to construct and operate the Project.

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20 Environmental Studies, Permitting, and Social or Community Impact
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This section provides an overview of the environmental, social, and permitting context of the Project. It describes the current baseline conditions and ongoing studies, outlines existing authorizations and future permitting requirements, and summarizes management plans for water, waste disposal, and environmental monitoring. Socio-economic baseline conditions, community engagement, and mine closure and reclamation planning are also addressed. The information presented is drawn primarily from the baseline and supporting studies have been completed by Tierralta, on behalf of NPM, in 2024.

The Carangas Project is located in the Carangas region in the western portion of the Department of Oruro, Bolivia, approximately 190 km northwest of the city of Oruro (population around 300,000), and within 50 km of the international border with Chile. Refer to Figure 5-1, Section 5, for the project general location plan.

The Project is situated in the high-altitude, arid volcanic landscape of the western Bolivian Altiplano with elevations ranging from approximately 3900 masl to 4100 masl. The region features scarce water resources, sparse high-elevation drought-resistant Andean vegetation typical of the semi-arid Altiplano steppe, and high-altitude fauna, and pastoral fauna dominated by domestic camelids. Some small farming and grazing communities (pueblos) are scattered throughout the region, the closest being Carangas, situated between the two prominent hills (the West Dome and the East Dome) on the Property. The official population of Carangas pueblos within the area of influence for the project reaches 1,130 inhabitants according to the 2024 national census (INE, 2024).

The Project contemplates the development of an open pit precious metal and base metal surface mine, together with a central processing plant, ore stockpiles, tailings storage facility using thickened slurry technology, waste rock management areas, and supporting infrastructure including access roads, power supply, fuel storage, explosive storage area, water supply and management systems, and ancillary facilities (refer to Figure 18-1 for general arrangement of proposed site infrastructure).

The proposed mine life consists of two years of construction followed by sixteen years of mine operations and three years of post-pit mining stockpile rehandle to the mill, followed by a closure and post-closure phase. The average ore processing rate will be 8.0 Mt/a (22 kt/d), increasing to 16.0 Mt/a (44 kt/d) in Year 6 of the project. Waste rock produced during the mine life is estimated to total approximately 363 Mt.

20.1 Socio-Environmental Considerations – Baseline and Supporting Studies

Select environmental baseline and supporting studies were completed by Tierralta on behalf of NPM. These studies were completed in 2024 and include work in the following areas: watershed characteristics; hydrology and meteorology; groundwater and surface water quality; wetlands; biodiversity of flora; and ambient air quality. Social and community studies consisted of a review of secondary data sources and focused on the following areas: population; migration and immigration; education; health; energy use, and economic activity. Summaries of the results of these studies are provided in the subsections below.

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The scope of these baseline studies is appropriate for the current design and permitting phase of the Project. As the Project moves into the feasibility design phase and environmental assessment/permitting process progresses, there are identified gaps and additional field, and desktop studies that will be required. The gaps currently include studies related to geochemistry and ARD/ML potential, noise, hydrogeology, faunal surveys, species at risk identification for flora and fauna, aquatic and fish studies, archaeological and cultural resource surveys, and focused studies on community land use within and adjacent to the Project area based on primary data sources.

20.1.1 Watershed Characteristics of the Project Area

The watershed in the vicinity of the Project is known as the Carangas micro-basin (or also know as a sub-basin). It is located within the Oruro Department and is one of the contributing sources of the Todos Santos River, which feeds the Sabaya River sub-basin. The hydrologic characteristics of the Carangas micro-basin are summarized in Table 20-1. The hydrologic network of the Carangas micro-basin and adjacent basins and divides are shown in Figure 20-1.

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Figure 20-1:      CarangasSub-Basin Hydrographic Map

Notes: Blue polylines denote water courses, streams, or rivers; Red polyline denotes boundary of Carangas sub-basin; Scale in meters.

Source: Tierralta, 2024

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Table 20-1:      Characteristicsof the Carangas Sub-Basin

No. Name Easting Northing Perimeter (km) Area (km²) Altitude (masl) Max. Altitude (masl) Min. Altitude (masl)
1 Carangas 537965.71 7905519.13 89.13 169.50 4276.06 4721 3870

Source: Tierralta, 2024

The Carangas micro-basin has an area of 169.50 km² and a perimeter of 89.13 km, delineated by the watershed divides. These values are essential for evaluating its size and shape, which, in turn, influence hydrological processes. With a Gravelius index of 1.93, the basin is elongated and not very compact. This means that water flow tends to concentrate in a single main channel, resulting in longer runoff times and a slower response to precipitation.

The basin has a relatively low slope, which promotes water infiltration and reduces surface runoff speed. Both the Drainage Density and River Frequency are moderate, indicating that the drainage system is adequate for evacuating water, except in cases of extreme rainfall. The concentration time from the furthest headwater to the outlet is estimated to be approximately five hours, which is consistent with the elongated shape and low slope of the basin.

In the case of the Carangas basin, the drainage density is 0.99 km/km², which classifies its drainage network with a low occurrence level. This suggests that the basin has a limited network of channels, which could reflect specific characteristics of the terrain, such as less permeable soils or gentle slopes.

In the case of the elevation difference, it was determined that the basin has an average of 0.85 km, which corresponds to the difference between the highest and lowest elevations throughout the basin; that is, from the highest point (the headwaters of the Kkotaña River) to the mouth of the Carangas River at the Todos Santos River.

The elevation difference of the main channel is 0.23 km and includes the Carangas, Vila Murata, and Kkotaña rivers. The maximum elevation of the basin reaches 4,550 masl, located at the headwaters of the Kkotaña River, while the average elevation is 4,280 masl. Finally, the mouth of the Carangas River is at an elevation of 4,051 masl.

For the classification of water bodies, a monitoring network was established consisting of nineteen (19) evaluation sites strategically distributed across different areas of the basin: five (5) in the headwaters (upstream), ten (10) in the mid-basin, two (2) in the lower basin (downstream), and an additional two (2) evaluation sites (control points) located at the mouth of the basin and another along the Todos Santos River. Refer to Table 20-2 and Figure 20-2 for details and locations of the surface water evaluation sites, relative to the Project area.

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Table 20-2:      SurfaceWater Evaluation Sites

Code Easting Northing Observation Sector
Upstream SW-KKO-10 537041 7908734 With Sample Kkotaña River upstream of Vila Murata River
SW-CTA-09 543158 7907414 With Sample Unnamed Creek, Taipihuano sector on the banks of Pucara Hill
S<W-CPU-08 543385 7904995 With Sample Jaillu Huani Creek, Pucara community sector
SW-JKH-15 542409 7908635 With Sample Jachcha Khalani Creek, originating from Kachi Pata Hill and Condoriri Hill
SW-QSNOM-16 541013 7908960 Without<br><br>sample (Dry) Unnamed River, upper sector on the banks of Wila Nasa Hill
Mid Zone SW-PBL-14 540473 7906468 With Sample Peña Blanca Creek, road to Pucara community
SW-CPU-13 542573 7905683 Without<br><br>sample (Dry) Unnamed Creek sector on the banks of Huakallani Hill
SW-QSNOM-07 542817 7908102 With Sample Jachcha Khalani Creek, Taipihuano sector on the banks of Condoriri Hill
SW-PCA-02 539496 7905937 With Sample Carangas River on the banks of Espíritu Santo Hill, approximately 65 m from Carangas Bridge
SW-VM-06 537975 7907406 With Sample Vila Murata River on the banks of Mamaniri Willkhi Hill
SW-VM-01 538126 7906526 With Sample Vila Murata River on the banks of Mamaniri Willkhi Hill and Espíritu Santo Hill
SW-QSNOM- 06 538040 7907536 Without <br><br>sample (Dry) Unnamed Creek, tributary of Vila Murata River upstream on the banks of Mamaniri Willkhi Hill
SW-QSNOM-17 540661 7908054 Without<br><br>sample (Dry) Unnamed River, middle sector on the banks of Wila Nasa Hill
SW-MKH-18 540826 7900830 Without<br><br>sample (Dry) Marca Khaua Creek, originating from Churi Kkollu Hill and Torreni Hill
SW-FUN-19 540554 7904656 Without<br><br>sample (Dry) Fundición Creek, originating from Conserva Hill and Huakallani Hill
Downstream SW-CA-03 538710 7904619 With Sample Carangas River on the banks of Conserva Hill
SW-QMK-12 538666 7901245 Without<br><br>sample (Dry) Unnamed Creek on the banks of Conserva Hill and Chapi Kkollu Hill
Control points SW-PTS-04 537600 7898457 With Sample Todos Santos River, La Rivera sector, approximately 60 m from Todos Santos Bridge
SW-TS-05 540206 7897156 With Sample Todos Santos River, approximately 250 m from the road

Source: Tierralta, 2024

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Figure 20-2:      SurfaceWater Evaluation Sites

Notes : Con muestra : Evaluation site with water, sample collected ; Sin muestra : Evaluation site dry, no sample taken; scale in meters.

Source: Tierralta, 2024

20.1.2 Hydrology and Water Sources

The climate of the Western Cordillera is cool and dry, especially in the winter months. In the Carangas area, high temperatures range from 12.4°C in July to 19.2°C in October, and the low temperatures range from -3.5°C in July to 3.8°C in January. Rainfall in the area is sparse and ranges from 2 mm in June to 162 mm in January. Refer to Table 5-1 for monthly climate averages. The wet periods correspond to the months from December to March. On the other hand, the dry periods cover the months from May to August. Finally, regarding precipitation the transition periods are found in the months of April, September, October, and November. The climatic features show scarce precipitation; the climatic boundary indicates that evaporation exceeds annual precipitation.

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Flow measurements were made at multiple locations throughout the stream basin in October 2024 (during the dry season) at the established evaluation sites. It is important to note that six of the sites were reported as dry watercourses (absence of water), as shown in Table 20-3.

By summing the flow of the basin under study (upstream, mid-basin, and downstream), the accumulated flow is 0.029 (m³/s). At the control points, for code SWPTS-04, the flow is 0.8541 (m³/s), and for SW-TS-05, the flow is 0.5089 (m³/s).

Table 20-3:      FlowMeasurements

Upstream Mid-basin Downstream Control Points
Code SW- KK0- 10 SW- CTA- 09 SW- CPU- 08 SW- JKH- 15 SW- QSNOM- 16 SW- PBL- 14 SW- CPU- 13 SW- PCA- 02 SW- VM- 06 SW- VM- 01 SW- QSNOM- 06 SW- QSNOM- 17 SW- MKH- 18 SW- MKH- 19 SW- CA- 03 SW- MKH- 12 SW- PTS- 04 SW- TS- 05
Flow<br> (m^3^/s) 0.007 0.0002 0.0001 0.0001 - 0.003 0.0001 0.006 0.006 0.006 - - - - 0.0012 - 0.8541 0.05089

Note: - denotes dry watercourse with no observed flow. Source: Tierralta, 2024

Although there are springs (phuxus) and wetlands (juqhus) that are primarily used for human consumption, there are also springs (Wañaqu, Jamp’atiri, Sik’iri Qalaqullu, and Viscachani) and rivers (Carangas, La Rivera, Mantos, and Qutaña) that are used for irrigation in agricultural activities, particularly benefiting the communities near these watercourses. However, some of these are seasonal rivers, so there is a primary reliance on meteorological or atmospheric sources, mainly from natural rainfall (MDRyT, 2012) and Yana (2017). The availability of water reaches its peak between the months of November and March, with the highest intensity in January and February, gradually decreasing during the other months, sometimes even leaving some rivers without surface water. Water shortages have historically occurred in the months of September, October, and November. A preliminary list of identified water sources was compiled as part of baseline studies.

In the territory of Carangas, water resources are both surface and underground. Among the main rivers that pass through the different communities of Carangas Marka are the Carangas River, Qutaña River, and Mantos River. Additionally, there are seasonal rivers that specifically emerge during the rainy season (MDRyT/VT, 2021).

The water used by families for consumption in their homes mainly comes from springs and ditches (67%), while 14% of families have access to a public water fountain, and 11% have piped water, which primarily benefits the families in the town center of Carangas. Additionally, households consider drought periods and the hydrological cycle when adopting alternative strategies.

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The importance of water resources is determined by the uses made of this resource. On one hand, as is widely known, water is essential for human consumption. On the other hand, water is important for the raising of domestic animals and, in communities with micro-irrigation systems, it is also used for crop irrigation.

20.1.3 Groundwater and Surface Water Quality
20.1.3.1 Groundwater Quality
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In October 2024, Tierralta collected groundwater quality samples from three installed wells in the study area as described in Table 20-4.

Table 20-4:      GroundwaterEvaluation Sites

Code Easting Northing Sector
Groundwater PP-CRG-01 540781 7908063 Water sample taken from Well in the Cargadero Sector, belonging to Ayllu Tunari, on the shores of Cerro Wila Nasa.
PP-SA-02 540951 7906860 Water sample taken from Well at the foothills of Cerro San Antonio.
PP-ISIW 538555 7904734 Water sample taken from Pisaqueri-Isiwaña, 200 meters from the Carangas River.

Source: Tierralta, 2024

The groundwater quality test results are illustrated in Table 20-5, and compared to permissible limits established by the Water Pollution Regulations (Reglamento en Materia de Contaminación Hídrica, RMCH), the regulatory body governing water pollution in Bolivia, where applicable. The Class A and Class D water quality criteria are provided for comparison purposes: Class A being the best water quality (pristine waters requiring little or no prior treatment for domestic or agricultural use), while Class D waters are of poor quality and to be avoided for domestic or agricultural use, but may be suitable for industrial use.

Table 20-5:      GroundwaterQuality Test Results

Parameters Units Oct-24 Oct-24 Oct-24 RMCH
PP-CRG- 01 PP-SA-02 PP-ISIW- 03 A D Unit
pH 7.8 7.8 6.2 6 to 8.5 6 to 9 -
Temperature °C 14.8 17.1 15.5 3 +/- 3 +/- °C
Conductivity μS/cm 266 394 5160 - - -
Settle Solids ml/L 2 <0.1 <0.1 <10 <100 mg/L
Dissolved Solids mg/L 170 255 3300 1000 1500 mg/L
Suspended Solids mg/L 224 <1 <1 - - -
Floating Solids Presence Absence Absence Absence <ret.malla 1mm^2^ -
Color (436 nm) mg/L 23 2 8 <10 <200 -
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Parameters Units Oct-24 Oct-24 Oct-24 RMCH
--- --- --- --- --- --- --- ---
PP-CRG- 01 PP-SA-02 PP-ISIW- 03 A D Unit
Turbidity NTU 434 5.08 6.59 <10 <200-10000 UNT
Dissolved Oxygen mg/L 6.4 5.6 1 >80% sat >50% sat mg/L
Redox Potential mV 111 89 142 - - -
Hardness mg/L 57.4 96 285.7 - - -
Chemical Oxygen Demand (COD) mg/L 12 20 17 <5 <60 mg/L
DBO5 mg/L 5 9 10 <2 <30 mg/L
Thermotolerant Coliforms UFC/100 mL 0 0 0 - - -
Total Coliforms UFC/100 mL 0 0 0 - - -
Oils and Fats mg/L <0.3 <0.3 <0.3 Absence 1 mg/L
Alkalinity mg/L 60 95 1240 - - -
Chromium VI mg/L 0.012 0.014 <0.005 0.05 c. Cr 0.05 c. Cr+6 mg/L
Ammonia mg/L 0.1 0.12 0.25 0.05 c. NH3 4 c. NH3 mg/L
Carbonates mg/L <0.1 11 <0.1 - - -
Bicarbonates mg/L 37 36.5 551 - - -
Chloride mg/L 22.7 31.3 1121.5 250 c. CI 500 c. Cl mg/L
Total Cyanide mg/L <0.05 <0.05 <0.05 0.02 0.2 mg/L
Total Phosphorus mg/L 0.14 0.06 0.13 - - -
Phosphate mg/L 0.18 0.08 0.17 0.4 c. 1.0 c. mg/L
Fluoride mg/L <0.01 <0.01 <0.01 0.6–1.7 c. F 0.6–1.7 c. F mg/L
Nitrates mg/L 19.33 7.17 2.13 20.0 c. NO3 50.0 c. NO3 mg/L
Nitrites mg/L 0.04 <0.02 <0.02 <1.0 c. N 1.0 c. N mg/L
Sulfates mg/L 32.7 51.9 84.9 300 c. SO4 400 c. SO4 mg/L
Sulfur mg/L 0.005 0.008 0.006 0.1 1 mg/L
Detergents mg/L 0.047 0.001 0.002 0.5 0.5 mg/L
Phenols mg/L 0.095 0.122 0.103 1 c. C6H5OH 10 c. C6H5OH ug/L

Notes: Red highlighted cells indicate that Class D water quality criterion is exceeded. Yellow highlighted cells indicate that Class A water quality criterion is exceeded or may be exceeded (based on analytical detection limit). Source: Tierralta, 2024

As shown in Table 20-6, all three samples exceeded the Class A criteria for biological and chemical oxygen demand and one sample displays elevated dissolved solids. Total cyanide and oils and fats are highlighted as exceeding Class A criteria due to elevated analytical detection limits.

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20.1.3.2 Surface Water Quality
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As discussed in Section 20.1.1 and presented in Table 20-2 and Figure 20-1, a monitoring network was established consisting upstream headwater sites (5), mid-basin sites (10), downstream lower basin sites (2), plus reference/control sites (2).

Out of the total of 19 evaluated sites, seven sites were reported as dry, and a total of 12 surface water quality samples were collected. The sample results were compared with the permissible limits established by the RMCH referencing Class A and Class D criteria. These are summarized in Table 20-6, Table 20-7, and Table 20-8 which present the analytical results for upstream, midbasin, and downstream/control points locations, respectively.

For the most part, surface water quality is good throughout the Carangas micro-basin meeting Class A criteria for most parameters. For the 25 parameters considered for laboratory analysis, the values for ammonia, BOD5, and COD consistently exceed the permissible limits for Class A, but comply with the permissible limits for Class D. Specifically, for the pH parameter, several samples report values above pH 8 and one sample reports a turbidity exceedance. Dry seasonal conditions resulted in seven sites not being sampled.

Table 20-6:      SurfaceWater Quality (upstream)

Parameters Units Upstream RMCH
SW-KKO-10 SW-CTA-09 SW-CPU-08 SW-JKH-15 A D Unit
pH 9.0 9.1 7.9 8.7 6–8.5 6–9 -
Temperature °C 14.2 15.7 11.8 14.2 3 +/- 3 +/- °C
Conductivity µµS/cm 371.0 344 555 670 - - -
Dissolved Solids mg/L 238.0 220 355 429 1000 1500 mg/L
Suspended Solids mg/L 1.0 1 64 1 - - -
Turbidity NTU 2.60 1.77 4.94 1.71 <10 200-10000 UNT
Dissolved Oxygen mg/L 6.5 9.4 7.6 10.2 >80% >50% mg/L
Flow Rate m^3^/s 0.007 0.0002 0.0001 0.0001 - - -
Ammonia mg/L 1.65 1.98 1.12 2.39 0.05 c. NH3 4 c. NH3 mg/L
Total Phosphorus mg/L 0.07 0.09 0.08 0.04 - - -
Mercury mg/L <0.001 <0.001 <0.001 <0.001 0.001 0.001 mg/L
Phosphate mg/L 0.11 <0.04 0.24 <0.04 0.40 1.00 mg/L
Nitrates mg/L 3.24 1.64 1.52 5.82 20 50 mg/L
BOD5 mg/L 7 6 5 8 <2 <30 mg/L
COD mg/L 22 20 20 29 <5 <60 mg/L
Boron mg/L 0.95 1.45 1.97 1.21 1.0 c. B 1.0 c. B 1.0 c. B
Fecal Coliforms UFC/100 m 0 0 3 0 - - -
Aluminum mg/L <0.1 <0.1 <0.1 <0.1 0.2 c. Al 1.0 c. Al mg/L
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Parameters Units Upstream RMCH
--- --- --- --- --- --- --- --- ---
SW-KKO-10 SW-CTA-09 SW-CPU-08 SW-JKH-15 A D Unit
Arsenic mg/L 0.03 0.02 0.02 0.02 0.05 c. As 0.1 c. As mg/L
Cadmium mg/L <0.012 <0.012 <0.012 <0.012 0.005 0.005 mg/L
Copper mg/L <0.012 <0.012 <0.012 <0.012 0.05 c. Cu 1.0 c. Cu mg/L
Iron mg/L 0.131 <0.018 <0.018 <0.018 0.3 c. Fe 1.0 c. Fe mg/L
Manganese mg/L <0.009 <0.009 <0.009 <0.009 0.5 c. Mn 1.0 c. Mn mg/L
Lead mg/L <0.109 <0.109 <0.109 <0.109 0.05 c. Pb 0.1 c. Pb mg/L
Zinc mg/L <0.012 <0.012 <0.012 <0.012 0.2 c. Zn 5.0 c. Zn mg/L

Notes: Red highlighted cells indicate that Class D water quality criterion is exceeded. Yellow highlighted cells indicate that Class A water quality criterion is exceeded. One sampling site (QSNOM-16) was reported as dry and was therefore not sampled. Source: Tierralta, 2024

Table 20-7:      SurfaceWater Quality (Middle)

Parameter Units Mid-Basin Zone RMCH
SW-PBL-14 SW-QSNOM-07 SW-PCA-02 SW-VM-06 SW-VM-01 A D Unit
pH 9 8 8.5 8.5 8.3 6–8.5 6–9 -
Temperature °C 13.5 13.3 8.6 13.2 12.2 3 +/- 3 +/- °C
Conductivity µµS/cm 330 660 422 388 415
Dissolved Solids mg/L 211 422 270 248 266 1000 1500 mg/L
Suspended Solids mg/L <1 <1 136 <1 <1 - - -
Turbidity NTU 1.96 1.87 13.3 2.48 1.88 <10 200-10000 UNT
Dissolved Oxygen mg/L 7.9 7.8 7.6 6.6 6.4 >80% >50% mg/L
Flow Rate m^3^/s 0.003 0.0001 0.006 0.006 0.006 - - -
Ammonia mg/L 1.65 1.23 1.94 1.66 1.28 0.05 c. NH3 4 c. NH3 mg/L
Total Phosphorus mg/L 0.11 0.06 0.08 0.08 0.07 - - -
Mercury mg/L <0.001 <0.001 <0.001 <0.001 <0.001 0.001 0.001 mg/L
Phosphate mg/L 0.16 <0.04 0.22 0.04 0.06 0,4 1 mg/L
Nitrates mg/L 7.54 4.47 4.96 3.49 2.5 20 50 mg/L
BOD5 mg/L 6 10 8 8 6 <2 <30 mg/L
COD mg/L 18 30 30 24 19 <5 <60 mg/L
Boron mg/L 1.97 0.87 1.02 1.16 1.18 1.0 c. B 1.0 c. B 1.0 c. B
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Parameter Units Mid-Basin Zone RMCH
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SW-PBL-14 SW-QSNOM-07 SW-PCA-02 SW-VM-06 SW-VM-01 A D Unit
Fecal Coliforms UFC/100 m 1 0 15 0 5 - - -
Aluminum mg/L <0.1 <0.1 <0.1 <0.1 <0.1 0.2 c. Al 1.0 c. Al mg/L
Arsenic mg/L 0.03 0.02 0.04 0.04 0.03 0.05 c. As 0.1 c. As mg/L
Cadmium mg/L <0.012 <0.012 <0.012 <0.012 <0.012 0.005 0.005 mg/L
Copper mg/L <0.012 <0.012 <0.012 <0.012 <0.012 0.05 c. Cu 1.0 c. Cu mg/L
Iron mg/L <0.018 <0.018 <0.018 0.069 0.037 0.3 c. Fe 1.0 c. Fe mg/L
Manganese mg/L <0.009 <0.009 <0.009 <0.009 <0.009 0.5 c. Mn 1.0 c. Mn mg/L
Lead mg/L <0.109 <0.109 <0.109 <0.109 <0.109 0.05 c. Pb 0.1 c. Pb mg/L
Zinc mg/L <0.012 <0.012 <0.012 <0.012 <0.012 0.2 c. Zn 5.0 c. Zn mg/L

Notes: Yellow highlighted cells indicate that Class A water quality criterion is exceeded. Five sampling sites (SW-CPU-13, SW-QSNOM-06, SW-QSNOM-17, SW-MKH-18, and SW-MKH-19) were reported as dry and were therefore not sampled. Source: Tierralta, 2024

Table 20-8:      SurfaceWater Quality – Downstream and Control Points

Parameters Units Downstream Control Point RMCH
SW-CA-03 SW-PTS-04 SW-TS-05 A D Unit
pH 7.9 8.8 8.8 6–8.5 6–9 -
Temperature °C 3.1 14 13.7 3 +/- 3 +/- °C
Conductivity µµS/cm 1158 746 873 - - -
Dissolved Solids mg/L 742 479 558 1000 1500 mg/L
Suspended Solids mg/L <1 1 2 - - -
Turbidity NTU 2.02 3.7 3.82 <10 200-10000 UNT
Dissolved Oxygen mg/L 9.3 7.7 7.1 >80% >50% mg/L
Flow Rate m^3^/s 0.0012 0.8541 0.5089 - - -
Ammonia mg/L 1.71 2.17 1.64 0.05 c. NH3 4 c. NH3 mg/L
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Parameters Units Downstream Control Point RMCH
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SW-CA-03 SW-PTS-04 SW-TS-05 A D Unit
Total Phosphorus mg/L 0.08 0.05 0.19 - -
Mercury mg/L <0.001 <0.001 <0.001 0.001 0.001 mg/L
Phosphate mg/L 0.15 0.07 0.14 0.4 1 mg/L
Nitrates mg/L 6.8 5.33 4.96 20 50 mg/L
BOD5 mg/L 7 <5 8.6 <2 <30 mg/L
COD mg/L 19 19 22 <5 <60 mg/L
Boron mg/L 3.41 1.53 1.87 1.0 c. B 1.0 c. B 1.0 c. B
Fecal Coliforms UFC/100 m 156 0 0 - - -
Aluminum mg/L <0.1 <0.1 <0.1 0.2 c. Al 1.0 c. Al mg/L
Arsenic mg/L 0.07 0.21 0.24 0.05 c. As 0.1 c. As mg/L
Cadmium mg/L <0.012 <0.012 <0.012 0.005 0.005 mg/L
Copper mg/L <0.012 0.019 <0.012 0.05 c. Cu 1.0 c. Cu mg/L
Iron mg/L <0.018 0.128 0.099 0.3 c. Fe 1.0 c. Fe mg/L
Manganese mg/L 0.095 <0.009 <0.009 0.5 c. Mn 1.0 c. Mn mg/L
Lead mg/L <0.109 <0.109 <0.109 0.05 c. Pb 0.1 c. Pb mg/L
Zinc mg/L 0.136 <0.012 <0.012 0.2 c. Zn 5.0 c. Zn mg/L

Note: Yellow highlighted cells indicate that Class A water quality criteria is exceeded. One sampling site (9SW-QMK-120) was reported as dry and was therefore not sampled. Source: Tierralta, 2024

20.1.4 Biodiversity Assessment

The diversity and floristic composition were assessed at 10 evaluation sites using the transect method, with an evaluation area of 50 x 3 meters. In general, a total of nine families, 15 genera, and 18 species were recorded (plant quantity, coverage, point line) for the slopes, as shown in Table 20-9.

Table 20-9:      Numberof Abiotic and Biotic Records of Flora

Taxonomic Levels Number of Flora Records
Number of Physical (Abiotic) Records 412
Number of Biotic Records 536
Species 18
Genera 15
Families 9

Source: Tierralta, 2024

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A total of 18 species have been identified, belonging to nine families at the evaluated sites. The most representative families are Asteraceae, with species such as Baccharis boliviensis, Baccharis tola, Chersodoma jodopappa, Parastrephia lepidophylla, Parastrephia quadrangularis, followed by Poaceae and Cactaceae, each with three species, and Fabaceae with two species. The remaining five families are represented by only one species each, as shown in Table 20-10.

Table 20-10:      Shrublandand Grassland Species

Family Scientific Name Common Name
Asteraceae Baccharis boliviensis phiño thola
Baccharis tola ñaka thola
Chersodoma jodopappa oke thola
Parastrephia lepidophylla supu thola
Parastrephia quadrangularis kullko thola
Cactaceae Airampoa ayrampo ayrampo airampu
Cumulopuntia boliviana puskallu
Echinopsis longispina sankawi
Caryophyllaceae Pycnophyllum tetrastichum taxa taxa
Ephedraceae Ephedra rupestris acero kainlla
Fabaceae Adesmia horrida añawaya
Adesmia polyphylla kosko añawi
Poaceae Deyeuxia sp. deyeuxia
Festuca orthophylla paja blanca
Jarava sp. sicuya
Rosaceae Tetraglochin cristata kainlla
Solanaceae Fabiana densa tara thola
Verbenaceae Junellia seriphioides kanlla macho

Source: Tierralta, 2024

The study area presents slopes with resinous and spiny shrublands, rocky slopes with shrubs, plains dominated by grasslands, and shrublands with resinous and dwarf bushes.

Wetlands were observed with watercourses, where only Lobelia oligophylla was recorded, a plant known for growing in watercourses. Other species such as Plantago tubulosa and Baccharis acaulis adapt to water scarcity. Therefore, it is important to assess the dynamics of water availability to determine the wetland's condition, as wetland vegetation plays a crucial role in the conservation and maintenance of aquatic ecosystems.

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The floristic composition and species diversity exhibit considerable variation depending on water availability, where the presence or absence of species is significant. For the wetlands in the study area, a total of 11 families, 19 genera, and 21 species were recorded, as shown in Table 20-11.

Table 20-11:      Numberof Abiotic and Biotic Records of Flora in Wetlands

Taxonomic Levels Number of Flora Records
Number of Physical (Abiotic) Records 385
Number of Biotic Records 1115
Species 21
Genera 19
Families 11

Source: Tierralta, 2024

A total of 21 species have been identified, belonging to 11 families at the evaluated sites. The most representative are Cyperaceae with 5 species, one of which could not be identified, followed by Asteraceae with four species, Poaceae with three species, Juncaceae with two species, and the remaining seven families each with only one species, as shown in Table 20-12.

Table 20-12:      WetlandSpecies

Family Scientific Name
Apiaceae Lilaeopsis macloviana
Asteraceae Baccharis acaulis
Asteraceae Cotula mexicana
Asteraceae Hypochaeris taraxacoides
Asteraceae Werneria pygmaea
Campanulaceae Lobelia oligophylla
Caryophyllaceae Arenaria digyna
Cyperaceae Carex sp. 1
Cyperaceae Carex sp. 2
Cyperaceae Eleocharis albibracteata
Cyperaceae Phylloscirpus deserticola
Cyperaceae Sp.
Gentianaceae Gentiana sedifolia
Juncaceae Juncus stipulatus
Juncaceae Patosia clandestina
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Family Scientific Name
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Juncaginaceae Triglochin concinna
Plantaginaceae Plantago tubulosa
Poaceae Deyeuxia sp
Poaceae Distichlis humilis
Poaceae Festuca rigescens
Ranunculaceae Ranunculus breviscapus
Abiotic Water
Abiotic feces
Abiotic Llama feces
Abiotic Stuble/dry grass residue
Abiotic Rock
Abiotic Salt
Abiotic Soil

Source: Tierralta, 2024

20.1.5 Ecosystem Functions

The selected ecosystem evaluation sites correspond to lotic ecosystems, that is, sections of rivers where the movement of water was evident. Table 20-13 presents the descriptions and locations of the evaluated sites.

Three indices (IHF, QBR, and IPA) were used to evaluate and describe the ecosystems:

· The IHF index is the Índice de Hábitat Fluvial (Fluvial Habitat Index), and measures<br>the physical quality of a river channel to assess the suitability of the river to support aquatic life.
· The QBR index is the Qualitat del Bosc de Ribera (Riparian Forest Quality Index), and measures<br>the quality of vegetation along riparian zones to assess the health of a riverbank ecosystem.
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· The IPA index is the Índice de Presión Antropogénica (Anthropogenic Pressure<br>Index), and measures the intensity of human impacts on an ecosystem, commonly in rivers or watersheds.
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Table 20-13:      LoticEcosystem Evaluation Site List

No. Code Easting Northing Observation Sector
1 VA-01 538126 7906526 With Sample Vila Murata River on the shores of Mamaniri Willkhi Hill and Espíritu Santo Hill
2 VA-02 539423 7905937 With Sample Carangas River on the shores of Espíritu Santo Hill, approximately 65 m from the Carangas Bridge
3 VA-03 538710 7904619 With Sample Carangas River on the shores of Conserva Hil
4 VA-04 537600 7898457 With Sample Todos Santos River, La Rivera sector, approximately 60 m from the Todos<br><br>Santos Bridge
5 VA-05 540706 7897166 With Sample Todos Santos River, approximately 250 m from the highway
6 VA-06 538040 7907536 Dry river Unnamed Stream, tributary of Vila Murata River upstream on the shores of Mamaniri Willkhi Hill
7 VA-07 542817 7908102 With Sample Jachcha Khalani Stream, Taipihuano sector on the shores of Condoriri Hill
8 VA-08 543385 7904995 With Sample Jaillu Huani Stream, Pucara community sector
9 VA-09 543158 7907414 With Sample Unnamed Stream, Taipihuano sector on the shores of Pucara Hill
10 VA-10 538666 7901245 Dry river Unnamed Stream on the shores of Conserva Hill and Chapi Kkollu Hill
11 VA-11 537975 7907406 With Sample Vila Murata River on the shores of Mamaniri Willkhi Hill
12 VA-12 537041 7908734 With Sample Kkotaña River upstream from Vila Murata River
13 VA-13 542275 7908469 With Sample Jachcha Khalani Stream, originating from Kachi Pata Hill and Condoriri Hill
14 VA-14 541033 7908960 Dry river Unnamed River, upper sector on the shores of Wila Nasa Hill
15 VA-15 542573 7905683 Dry river Unnamed Stream, sector on the shores of Huakallani Hill
16 VA-16 540473 7906468 With Sample Peña Blanca Stream, road to Pucara community
17 VA-17 540661 7908054 Dry river Unnamed River, middle sector on the shores of Wila Nasa Hill
18 VA-18 540554 7904656 Dry river Fundición Stream, originating from Conserva Hill and Huakallani Hill
19 VA-19 540826 7900830 Dry river Marca Khaua Stream, originating from Churi Kkollu Hill and Torreni Hill

Source: Tierralta, 2024

Out of the 19 evaluated sites, ten were categorized as important, while two were classified with a moderate level of importance. This stratification reveals an ecosystem with limited capacity to provide water regulation services, whose functionality is primarily focused on supplying resources for local livestock and surrounding communities.

20.1.6 Ambient Air Quality

In 2024, ambient air quality testing was completed over seven sampling locations to measure TSP, PM10, CO, SO2, NOx, and CO2, a comparison of the results obtained with the values defined in the Bolivian Regulation on Atmospheric Pollution to determine the air quality in the study area. The concentrations were all below the permissible limits set by the Bolivian regulations. This sampling was done during the rainy season, and therefore, particularly for dust, one would expect a lower concentration. The company is expected to conduct sampling during the dry season as part of the sampling program.

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20.1.7 Community Assessments and Studies
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20.1.7.1 Population
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According to official data from the 2024 National Population and Housing Census, the population of the project's area of influence reaches 1,130 inhabitants (INE, 2024); women represent 49.81% and men 50.19%, distributed across four Ayllus and 23 communities in the Municipality. Finally, considering the study conducted by MDRyT/VT (2014), the household size is five members per family, with 46 remaining families living in the area and 150 resident families.

The population pyramid of the Municipality of Carangas shows a wide base in the age ranges of 0 to 24 years and a narrow peak as ages increase. This is typical of societies with demographic inefficiency (many births, many deaths, and low life expectancy at birth) and that have not yet entered the demographic transition and population maturity (low proportion of births, low proportion of deaths, and high life expectancy at birth).

For the study of population distribution, three analysis groups are considered: the first consisting of children and young people, the second of adults, and the third of older adults. In this context, the population of children and young people reaches 394 inhabitants; adults total 684 inhabitants, and finally, older adults amount to 52 inhabitants. Therefore, it was concluded that the largest population in Carangas falls between the ages of 15 and 64.

20.1.7.2 Migration and Immigration

The net internal migration rate for the period from 2001 to 2012 was -8.6 (INE, 2014). According to estimates made by Tierralta S.R.L for the period from 2012 to 2024, the migration rate reaches -6.3. Migration in the municipal context is relatively dynamic, with temporary and/or permanent movements.

The first type of migration forms part of the local life dynamics, where movement is planned according to the agricultural cycle; from September to April, most people engage in agricultural and livestock work. During the dry season, from May to August, they move to capital cities in search of work, and young people go to pursue higher education.

20.1.7.3 Population Projections

The population projection is an estimate for Carangas and can vary depending on several factors, such as changes in the population growth rate, migration, and other demographic factors. The population projections for Carangas are based on certain technical criteria, considering the population growth rate, correlation coefficient, and the least squares method.

For the analysis, the base year is 2024, using the official data from the National Census of Population and Housing (Census Nacional de Población y Vivienda, CNPV), conducted that year. Moreover, the methods best suited for population projection are the geometric and exponential methods, as the municipality of Carangas has a growth rate close to 7% (INE, 2012), and with technical adjustments, it reaches 6.4%, which is an unusual situation that generally does not occur in other demographic contexts, resulting in a demographic explosion. In the projected period from 2025 to 2040, with the geometric method, it varies from 1,158 to 1,677 inhabitants, and with the exponential method, from 1,163 to 1,692 inhabitants.

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20.1.7.4 Poverty Levels
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Considering the condition of unmet basic needs, 72.9% of the population in Carangas is in moderate poverty, 23.9% is in a situation of indigence, and 0.8% is in marginal poverty, as shown in Table 20-14. Consequently, for the entire population of the municipality, the percentage of poor population reaches 97.6%.

Table 20-14:      Populationby Condition of Unmet Basic Needs

Condition of Unmet Basic Needs Population Percentage Distribution
Total Population (Subject of Study) 838 100
Not Poor UBN 3 0.4
Threshold 17 2
Poor Moderate 611 72.9
Indigent 200 23.9
Marginal 7 0.8
Percentage of Poor Population 97.6 - -

Source: Tierralta, 2024

20.1.7.5 Education

Regarding educational aspects, 83% of the population ages 6 to 19 in Carangas attends a public educational unit, 6% attend a private or partnership unit, respectively. However, it can be noted that 11% of this age group does not attend any educational center.

Regarding the school dropout rate, it is worth noting that no dropouts were reported in 2022. Additionally, the number of students enrolled at the initial level is 7, at the primary level 18 students, and at the secondary level 12 students. Among other aspects, the school breakfast program reaches approximately 37 students, which is implemented in coordination with the school boards and with the financial contribution of the Autonomous Municipal Government of Carangas. In order to improve student learning, the quality of breakfast and lunch has been improved. In the morning, a varied breakfast is provided, and at midday, a lunch with soup and a main course, similarly varied throughout the week, with a special dish on Wednesdays (PTDI, 2021-2025).

According to data from the sistema de informacióneducativa (Educational Information System, SIE) (2022), the Municipality of Carangas has two educational units. The first refers to U.E. Eliodoro Villazón, located in the municipality's capital, and the other is located in the Mantos community, an associated school that is administratively dependent on the direction of the Núcleo de Todos Santos (neighboring municipality) rather than the Eliodoro Villazón direction, despite being within the jurisdiction of the Municipality of Carangas. In this regard, the levels of education offered include community family education at the initial level, community vocational primary education, and community vocational secondary education.

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Based on the document review and information from the SIE (2022), the condition and quality of the infrastructure available at the educational units in the Municipality of Carangas as fair. However, it is considered important to perform periodic maintenance on the infrastructure, in addition to constructing new spaces for the student population. Furthermore, there is a need to equip the laboratories and the computer room.

The Eliodoro Camacho educational unit has 12 rooms, including classrooms, bathrooms, storage rooms, and laboratory rooms. The Mantos educational unit has one room for the initial level. Among other resources, there are six teachers, one director, and a janitor as part of the administrative staff.

On the other hand, the SWOT analysis conducted by Tierralta concludes that the Eliodoro Villazón educational unit has many weaknesses concerning its furniture and equipment. There is a need for more furniture, as the addition of new teachers and students increases the demand for school desks. This is a recurring problem that traditionally arises, and in some cases, it is addressed with loans, with some furniture being over 30 to 40 years old. Among other aspects, the physics and chemistry laboratory is another urgent need for secondary students; they lack equipment. Finally, no less important is the need for equipment in the music and physical education areas.

20.1.7.6 Health

Based on the Ministry of Health (2022), the municipality of Carangas is part of the Western Health Network and has a primary healthcare facility in the area, which shares the same name. The road connection between the municipality's capital and the communities is a dirt road, which becomes impassable during the rainy season, causing accessibility problems for the locals.

Regarding health aspects, the majority of the population (35%) in Carangas seeks care at a public healthcare facility. Additionally, 47% of the population relies on home remedies and/or visits from a traditional doctor. Furthermore, 15% of the population self-medicates or goes to pharmacies.

Based on the Ministry of Health (2022), the infrastructure and equipment of the health post in the municipality of Carangas are in fair condition. However, it is important to develop actions to strengthen health services and prevent diseases.

The health facility in Carangas is a first-level facility, with eleven rooms differentiated by specialty, as described in Table 20-15.

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Table 20-15:      Infrastructureof the Carangas Health Facility

Type of Health Facility Complies with Regulations Rooms Current Condition of Health Facility
Good Fair Poor
1st Level No Juana Azurduy Bonus Program - X -
General Medicine and Adolescent Care - X -
Telemedicine Program - X -
Pharmacy - X -
DOTS Area - X -
Nursing - X -
Expanded Immunization Program - X -
Delivery Room - X -
Maternity Room - X -
Meeting Room - X -
Health Personnel Rooms - X -

Source: Tierralta, 2024

The health facility in Carangas is staffed by one doctor and one nursing assistant. Additionally, there is temporary staff, including a general doctor, dentist, and ambulance driver. According to PTDI (2021-2025), the specialties offered are general medicine, nursing, and dentistry, with the addition of the Juana Azurduy Bonus program. Regarding the dental specialty, there are no dedicated rooms, and care is provided in a mobile dental unit.

Health coverage in the municipality of Carangas reaches approximately 60% of the population. The non-coverage for the remaining 40% of the population is due to several existing factors, such as the poor condition of roads, the distance to the health post from the communities, and the lack of available medications for medical care.

20.1.7.7 Energy Use

A large majority of families (86%) use firewood as a cooking fuel, 12% use gas in cylinders, and 2% of families use manure or dung. However, the dependence on firewood as an energy source can lead to health and environmental issues, as burning firewood can produce smoke and pollutants. Nevertheless, the municipality of Carangas has significant potential for renewable energy generation, such as solar and wind energy, which could help reduce the dependence on firewood and improve access to energy.

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20.1.7.8 Economic Activity
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According to the study conducted by MDRyT (2012) and Yana (2017), regarding the economic activities of the Carangas population, the following conclusions are made, in order of importance:

· Approximately 64% are engaged in agriculture and livestock
· About 16% are involved in camelid livestock farming
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· 14% are involved in livestock, commerce, and transportation
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· 6% are engaged in other activities such as livestock, construction, blacksmithing, and others.
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A total of 78% of the population are self-employed workers, while 9% are workers or employees dependent on the public and/or private sector. Up to 8% of the economically active population engaged in unspecified activities. The remaining percentage of the working population.

20.2 Environmental Management and Monitoring

Environmental monitoring and management for the Project will be designed to comply with Bolivian environmental legislation and to meet international good practice guidelines, including IFC Performance Standards. The monitoring program is intended to detect changes in environmental quality, verify the effectiveness of mitigation measures, and ensure compliance with environmental permit conditions.

20.2.1 Baseline Considerations and Derisking

The baseline studies completed to date for the Carangas Project provide a preliminary foundation of environmental and social information that can be built upon and expanded as needed. Hydrological and water quality assessments demonstrate the sensitive nature of surface water sources and the need to identify the presence of reliable perennial and intermittent streams and groundwater sources that can be incorporated in water management and diversion strategies. Geochemical studies are yet to be initiated and need to confirm acid generating and neutralizing capacity and the methodologies to be considered for managing potentially acid-generating mine waste through operations and closure. Ecological surveys have identified lotic ecosystems and wetlands that support diverse flora that will form the basis for future biodiversity management and restoration programs. Socio-economic studies of the Carangas community confirm rural land use dominated by agriculture and livestock farming; therefore, impacts to those traditional activities and to the community generally will need to be minimized by means of ongoing engagement and continued and ongoing environmental and social management practices and measures.

The baseline findings and supporting studies are used to help derisk the project by optimizing the mine layout, waste and tailings facility siting, and water management plans. Mitigation measures are being developed in line with Bolivian regulatory requirements, IFC Performance Standards, and community needs and will be refined through ongoing environmental and social management planning.

20.2.2 Water Supply and Management

Water supply requirements and management for the Project are presented in Sections 17.4.2 and 18.9 of this report. Currently it is proposed that raw water will be sourced through drilling in the project vicinity upstream of the open pit. In addition, alternative water sources are being identified and assessed.

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During mining operations noncontact water will be diverted away from the site to the greatest practical extent while contact water will be collected, potentially reused, and treated before being discharged to the environment. Additional work is required to evaluate the pit dewatering requirements, and the potential need to treat water from the dewatering wells prior to discharge to the environment. Noncontact water or treated contact water will be discharged to Carangas Creek. It is noted that assumed pit inflows and pit dewatering have not been considered in the context of the current study. It is also noted that a surface water diversion will be required around the proposed open pit.

20.2.3 Environmental Management and Monitoring Plans

Environmental management and monitoring for mining projects in Bolivia is regulated by the MMAyA, through national and departmental environmental authorities, under the General Environmental Law. Environmental monitoring will occur throughout the mine life, as indicated by national, departmental, and local regulations. Further post-closure environmental monitoring will align with the requirements for mine closure in Bolivia.

A comprehensive system of socio-environmental monitoring and management plans will be developed based on applicable regulatory requirements and considering community concerns and international best practices. Proposed construction and operations at the Project site will involve a range of activities focused on several localized areas where infrastructure is planned as described in the previous sections. To manage the Project, a suite of management and monitoring plans focussed on identified valued environmental and social components will be developed that will ensure the Project operates in a manner that protects Project personnel, the environment, and local communities, while promoting overall positive impacts. Plans for waste rock and tailings management are discussed in more detail in Section 18. The proposed conceptual methods of waste rock and tailings management are currently considered by regulators and standards associations as generally acceptable practices. Additional testwork will be required to determine the degree of potentially acid- generating (PAG) and non-acid generation (NAG).

A preliminary list of anticipated management and monitoring plans, based on current understanding of proposed Project infrastructure and activities, are listed below:

· Mine Emergency Response Plan
· Occupational Health and Safety Management Plan
· Spill Prevention and Contingency Plan
· Community Effects Monitoring and Management Plan
· Cultural Heritage Resources Protection Plan
· Terrestrial Environment Mitigation and Monitoring Plan
· Lotic Ecosystems and Wetlands Management Plan
· Erosion and Sediment Control Plan
· Fugitive Dust Control and Monitoring Plan
· Air Quality and Noise Abatement Management Plan
· Metal Leaching/Acid Rock Drainage Management Plan
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· Sampling Program QA/QC Plan
· Construction Environmental Management Plan
· Chemicals and Materials Storage, Transfer and Handling Plan
· Waste (Refuse and Emissions) Management Plan
· Hazardous Materials Management Plan
· Preliminary Reclamation and Closure Plan
· Freshwater Supply, Sewage and Wastewater Management Plan
· Soil Management Plan
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20.3 Permitting Considerations
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The Property consists of two mining areas, Granville and Colapso, granted by Autoridad Jurisdiccional Administrativa Minera (Mining Administrative Jurisdictional Authority, AJAM), the Bolivian mining authority. The two mining areas currently advancing through the transition process from Prospecting and Exploration Licenses (PELs) to Administrative Mining Contracts (AMC) in accordance with Bolivian mining regulations.

During 2025, NPM initiated and advanced this regulatory transition process. Work and Investment Plans for the Granville and Colapso mining areas were prepared and submitted to the relevant authorities and subsequently reviewed with SERGEOMIN and AJAM as part of the AMC evaluation and approval process.

Upon approval of the AMCs by AJAM, the Company intends to apply for environmental categorization and initiate preparation of the Environmental Impact Assessment (EEIA) required by future project environmental permitting for mine development.

20.3.1 Environmental Permits

In Bolivia, an environmental license is required before starting mining or exploration activities. The license is legally binding and may be suspended or revoked for non-compliance.

All projects are screened and assigned to one of four environmental categories, which determines the level of study and approval needed. These categories are illustrated in Table 20-16. Based on the scope of the Project, it is likely to fall under Category 1.

Table 20-16:      EnvironmentalLicensing System Categorization

Category Impact level Requirement Typical mining examples
Category 1 High impact Full Analytical Integral EIA (EEIA) Large-scale mining, processing plants
Category 2 Moderate impact Specific EIA/partial EEIA Medium operations, some exploration with disturbance
Category 3 Low impact Environmental Management Plan (EMP) Early-stage exploration, small works
Category 4 Minimal/negligible Simple registration (no EEIA) Desk studies, minimal disturbance

The permitting process consists of six main steps:

· Categorization Request - Submit project information to determine project category
· Preparation of Studies - Prepare the EIA or EMP depending on the project category
--- ---
· Submission to the Environmental Authority - Ministry of Environment (MMAyA) or departmental authority
--- ---
· Technical Review -Includes environmental and social evaluation as required
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· Approval and issuance of Environmental License
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· Post-approval obligations -Monitoring, reporting, auditing.
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20.3.2 Mining Permits
--- ---

The application process for the ELs to AMCs conversion started in 2025 and all required documents have been submitted to the Bolivia’s Ministry of Mining and Metallurgy since then. With the recently completed prior consultation, the Company expects that the remaining administrative work and legislative approval of the conversion of the ELs to AMCs could be completed within six months or less.

Once the AMCs are obtained, the Company will start the application to obtain its environmental categorization as a proposed open pit operation from Bolivia’s Ministry of Environment and Water, formally commencing the EEIA process. Work will also commence on gathering baseline environmental and social data as well as other associated technical work to fulfill the requirements of the EEIA. It is expected that this work will be completed by the end of 2027.

20.4 Social Considerations
20.4.1 Framework agreement with the Community
--- ---

The Carangas Project is located within the TerritorioIndígena Originario Campesino (Native Indigenous Peasant Territory, TIOC) of Carangas. NPM has adopted a proactive and progressive engagement strategy focused on long-term collaboration, transparency, and responsible development in accordance with Bolivian regulations and international mining practices.

Throughout 2024, 2025, and early 2026, NPM implemented an extensive community engagement program designed to strengthen relationships with local stakeholders and facilitate the advancement of the Project. Engagement activities included formal and informal meetings with community members and authorities, technical workshops, educational outreach programs, stakeholder mapping, sociocultural baseline studies, and communication campaigns tailored to local customs and traditions. Additional support initiatives included veterinary assistance and technical support programs for local camelid livestock producers, as well as participation in civic, social, and cultural activities within the community.

On January 26, 2026, NPM and the Carangas Community executed a formal Framework Agreement regulating the relationship between the Project and the Community. This agreement represents the most significant advancement achieved to date in obtaining the social permits required for project development and establishes the principal governance instrument currently in force between both parties.

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20.4.2 Prior Consultation Process
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The Prior Consultation process prior to the issuance of Mining Administrative Contracts, regulated by Bolivian mining legislation, has also progressed substantially. During 2025, AJAM completed the identification of consultation stakeholders through field verification activities and preparation of the corresponding technical reports. The formal Prior Consultation process is currently scheduled for July 2026.

The application process for the ELs to AMCs conversion started in 2025 and all required documents have been submitted to the Bolivia’s Ministry of Mining and Metallurgy since then. With the recently completed prior consultation, the Company expects that the remaining administrative work and legislative approval of the conversion of the ELs to AMCs could take up to six months.

20.4.3 Collaborative Work with Community

The Project currently anticipates that relocation of the existing Carangas community may be required due to the proximity of the mineral deposit and planned mining infrastructures to the current Carangas village. In preparation for this process, NPM has worked collaboratively with community representatives to identify potential locations for the future townsite. Architectural concepts, 3D visualizations, and physical models were prepared to support informed community decision-making. The community is expected to reach a preliminary decision regarding the preferred relocation site on June 15, 2026. Preliminary property valuation studies, socioeconomic baseline investigations, and conceptual resettlement planning have also been initiated with the support of specialized consultants.

Social and cultural considerations have already positively influenced project design optimization. In response to community concerns, the original TSF location was reassessed, and Klohn Crippen Berger (KCB) was retained to evaluate alternative locations. Following technical, environmental, and social evaluations, the “Cargadero” area north of the Project was selected as the preferred alternative location for TSF. This modification significantly reduced potential social conflict and improved overall project viability. Ongoing project planning also considers preservation of cultural heritage features, including the historic colonial church, as well as community territorial organization and future urban planning requirements.

NPM continues to recognize the importance of ongoing environmental and social management throughout future project development. Key social considerations include livelihood transition, community relocation, preservation of cultural infrastructure, health and safety, and long-term social cohesion.

Overall, the Carangas Project has achieved substantial progress in community engagement, permitting advancement, and social governance. The execution of the Framework Agreement, advancement of the Prior Consultation process, and incorporation of community feedback into project design demonstrate NPM’s commitment to responsible mining development and collaborative stakeholder engagement. These accomplishments significantly strengthen the Project’s long-term feasibility and social stability.

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20.5 Closure and Reclamation Planning
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20.5.1 Closure and Reclamation Plans
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Closure of the Project will be primarily regulated by the Viceministerio de Medio Ambiente (Deputy Minister of the Environment), part of MMAyA, AJAM, and local governments. The legal framework for mine closure in Bolivia is governed by three key instruments: Ley del Medio Ambiente No. 1333 (Environmental Law 1333), the Reglamento Ambiental para Actividades Mineras (Environmental Regulations for Mining Activities, RAAM), and the environmental licensing system (RPCA/EEIA). As part of the permitting stage, the Project must submit a Closure Plan along with the EEIA application. The Plan must include objectives, closure program details, rehabilitation measures, contaminant control measures, and post-closure monitoring. The Closure Plan will be prepared in parallel with mine facility designs, incorporate baseline information studies, and other operational and long-term planning efforts.

The project will be closed in two phases: the active closure phase and the passive post-closure phase. During the active closure phase, closure reclamation activities for the mining, process plant, infrastructure, and TSF will take place. Environmental monitoring is assumed to be conducted during the reclamation and post-closure phases. The requirement for water treatment will be further assessed based on water balance considerations and geochemistry source term studies to be completed as the Project progresses through advanced studies. Treatment of water utilizing passive treatment systems will be a consideration during the passive post-closure phase.

20.5.1.1 Mining

The preliminary closure strategy for the mining areas includes:

· Regrading stockpiles to overall slope angles
· Scarification of haul roads
--- ---
· Placement of 0.3 m of overburden on the tops and sides of the stockpiles and selected areas of the open<br>pit
--- ---
· Placement of 0.1 m of overburden on the haul roads
--- ---
· Placement of 0.3 m of topsoil on the tops and sides of stockpiles and selected areas of the open pits
--- ---
· Placement of 0.1 m of topsoil on the haul roads
--- ---
· Revegetation of the stockpiles and haul roads.
--- ---

Mining area closure activities are based on the use of the owner operated mine fleet for completing the activities listed above, except for the revegetation. These activities are planned to occur during the active closure phase.

20.5.1.2 Process Plant and On-Site Infrastructure

In preparation for closure of the mill building, there are four areas that will be addressed:

· Removing all machinery, equipment, building, and structures
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· Covering and revegetating concrete foundations
--- ---
· Disposing of scrap metal
--- ---
· Disposing of chemicals and reagents.
--- ---

These reclamation activities are planned to occur during the active closure phase.

20.5.1.3 Tailings Storage Facility

Upon closure, the following activities will be initiated with regards to the TSF:

· Placement of 2.0 m NAG waste rock material on the free crest and downstream slope of the tailings dam
· Placement of 0.3 m of topsoil on the free crest and downstream slope of the tailings dam
--- ---
· Revegetation of the free crest and downstream slope of the tailings dam
--- ---
· Redesign of closure spillway.
--- ---

These reclamation activities are planned to occur during the active closure phase.

20.5.1.4 Waste Rock Storage Facility

Upon closure, the following activities will be initiated with regards to the WRSF:

· Regrading WRSF to overall slope angles
· Placement of 0.3 m of topsoil on top and slopes of WRSF
--- ---
· Revegetation of the WRSF as required.
--- ---
20.5.1.5 Water Treatment
--- ---

Excess water in the water management pond will be treated, if required, and discharged to local waterways during the active closure period. Treatment of water utilizing passive treatment systems will be considered during the passive post-closure period.

20.5.1.6 Environmental

Environmental monitoring will be required during the active and passive closure periods. Monitoring requirements will be based on regulatory requirements and consultation with local communities.

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20.5.2 Closure Cost Estimates
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Closure liabilities for the Project have identified the physical reclamation activities detailed below. The estimated closure cost is detailed in Section 21 and assumes the following:

· All disturbed areas as a result of mining activity will be regraded to promote drainage
· All buildings on site will be removed and disposed of in an approved landfill
--- ---
· Water management structures will be removed, and dams breached
--- ---
· An annual operating and sustaining capital allowance will be allocated for site management and maintenance
--- ---
· An annual allowance will be allocated for environmental monitoring and regulatory compliance
--- ---
· An annual allowance will be allocated for long-term community considerations
--- ---
· The estimate assumes additive premiums and contingences to account for the level of detailed design, location<br>of the site, and other unknowns.
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21 Capital and Operating Costs
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21.1 Introduction
--- ---

The capital and operating costs described in this PEA are based on an open-pit mining operation with an 18.4-year mine life. The initial phase of the process plant is designed to treat 22,000 t/d of mineralized material, or 8.0 Mt/a, producing two concentrate products. The first expansion phase occurs in Year 5, and increases throughput to 44,000 t/d, or 16.0 Mt/a, expanding the concentrate circuit. The second expansion phase occurs in Year 9 and converts 22,000 t/d of concentrate production capacity to gold leaching and doré production capacity.

21.2 Capital Costs
21.2.1 Overview
--- ---

The capital cost estimate was developed in Q2 2026 to target a level of accuracy of -30% to +50%, which aligns with an Association for the Advancement of Cost Engineering (AACE) International Class 5 level estimate. The estimate includes mining, processing, on-site infrastructure, off-site infrastructure, project indirects, project delivery, owners’ costs, and provisions. The total initial capital costs for the Project are estimated at US$644.5 million, including US$55.3 million of capitalized operating costs, and US$110.3 million of contingency. The LOM expansion costs are estimated at US$422.7 and include both the concentrate plant and the gold plant. The sustaining costs are estimated at US$166.5 million, while the closure costs are estimated at US$149.8 million. The capital cost summary is presented in Table 21-1.

Table 21-1:      CapitalCost Summary

WBS Description Initial Capital Cost (US$M) Concentrate Plant Expansion Cost (US$M) Gold Plant Expansion Cost (US$M) Sustaining Cost (US$M) Total Cost (US$M)
1000 Mining $3.4 - - $14.6 $18.1
2000 Concentrate Plant $202.1 $202.1 - - $404.3
3000 Gold plant - - $42.9 - $42.9
4000 On-site Infrastructure $65.8 $20.3 $0.8 $151.9 $238.8
5000 Off-site Infrastructure $114.5 - - - $114.5
Total Direct Costs $385.9 $222.5 $43.8 $166.5 $818.6
6000 Project Preliminaries $31.7 $16.7 $3.3 - $51.7
7000 Project Delivery $45.6 $38.9 $7.9 - $92.4
8000 Owner’s Costs $15.7 - - - $15.7
Total Indirect Costs $93.0 $55.6 $11.2 - $159.8
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WBS Description Initial Capital Cost (US$M) Concentrate Plant Expansion Cost (US$M) Gold Plant Expansion Cost (US$M) Sustaining Cost (US$M) Total Cost (US$M)
--- --- --- --- --- --- ---
Total Direct + Indirect Costs $478.9 $278.0 $54.9 $166.5 $978.3
Contingency $110.3 $74.6 $15.1 - $200.0
Subtotal Capital Cost $589.2 $352.6 $70.0 $166.5 **$1,**178.3
Capitalized Opex $55.3 - - - $55.3
Closure Costs - - - $149.8 $149.8
Total Capital Cost $644.5 $352.6 $70.0 **$**316.3 **$1,**383.4

Note: Totals may not match due to rounding.

21.2.2 Basis of Estimate

The capital cost estimate was developed in Q2 2026 American dollars (US$). The estimate is based on budgetary quotations for equipment from recent advanced studies and execution projects, supplemented with Ausenco’s in-house database, and informed by Ausenco’s experience from similar operations in South America.

The following data were used as the basis of estimate:

· Mining schedules
· Engineering design by Ausenco, including but not limited to design criteria, equipment lists, and material<br>take-offs (MTOs)
--- ---
· Budgetary equipment quotes from similar recently completed advanced studies and execution projects, escalated<br>to Q2 2026
--- ---
· Additional data such as lang factors and indirect costs from similar recently completed studies and projects.
--- ---

The estimate also adhered to these parameters:

· No allowance was made for exchange rate fluctuations
· No additional escalation was added to the final estimate
--- ---
· No price growth or design allowance was included.
--- ---

An exchange rate of US$0.73 per C$1.00 was used to convert costs denominated in Canadian dollars to U.S. dollars.

21.2.3 Mine Capital Costs

Mine capital costs have been derived from contractor-supplied unit cost estimates for open pit mining operations applied to the Carangas mine plan and PEA production schedule.

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Pre-production mine operating costs (i.e., all mine operating costs incurred before mill start-up, covering 19.1 Mt of scheduled waste and oxide rock mining) are capitalized and included in the capital cost estimate. Pre-production pit operating costs include drill and blast, load and haul, support, and departmental overhead costs.

The mining contractor will provide the mobile fleet and maintenance facilities required for mining operations, with these costs embedded within the operating cost unit rates. Furthermore, all site development costs for mine operations, including activities such as clearing and grubbing, haul road construction, stockpile preparation, pit dewatering, and crushed rock production, are also capitalized.

The following mine operations infrastructure items are also capitalized:

· Site GPS (global positioning system)
· Mine survey gear and supplies
--- ---
· Radio communications systems
--- ---
· Geology, grade control, mine planning and management software licenses
--- ---
· Geotechnical instrumentation
--- ---
· Piping for pit dewatering
--- ---

Table 21-1 summarizes the Mine Area Capital Cost estimates for the Carangas PEA Project. It is the QP’s opinion that these estimates are reasonable for the location and planned mine development and can be used for a PEA.

Table 21-2:        MineArea Capital Cost Summary

Item US$M
Capitalized<br> Pre-Production Contractor Mining Operating Costs 52.7
Capitalized<br> Pre-Production Owner Mining Operating Costs 2.6
Site<br> Development Capital Costs 2.7
Mine<br> Operations Infrastructure Capital Costs 0.8
Total Initial Mining Capital 58.7

Note: Totals may not match due to rounding.

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21.2.4 Process Capital Costs
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The selection and sizing of process equipment requirements was based on process flowsheets and process design criteria as defined in Section 17. All major equipment was sized based on the process mass balance, as dictated by the process design criteria, to develop a mechanical equipment list (MEL). The MEL was then developed through recent budgetary quotations. The remaining value of the equipment list was developed through benchmarking against recent execution projects and advanced studies.

The process plant and infrastructure engineering design was developed at a conceptual level consistent with a PEA, allowing for the bulk material quantities (steel, concrete, piping, cables, instruments, etc.) to be derived for the major commodities using Lang factors. Plant earthworks costs were derived from MTOs.

There are no sustaining costs associated with the process plant.

The total initial capital costs for the process plant are US$202.1 million. The capital cost breakdown for the process plant is summarized in Table 21-3.

Table 21-3:        ProcessPlant Capital Cost Breakdown

WBS Description Initial Costs (US$M)
2100 Crushing $22.5
2200 Grinding $75.1
2300 Flotation &<br> Regrind $49.0
2400 Concentrate<br> Dewatering $10.3
2500 Tailings<br> Dewatering $6.2
2600 Reagents $15.5
2700 Water &<br> Air Services $23.6
2000 Concentrate Plant $202.1

Note: Totals may not match due to rounding.

21.2.5 Infrastructure Capital Costs
21.2.5.1 On-site infrastructure
--- ---

The on-site infrastructure costs consist of bulk earthworks, power switchyard and distribution, fuel storage, sewage, potable water, water management, infrastructure buildings, tailings facility and pipelines, permanent camp, and site services and mobile equipment. The total on-site infrastructure costs for the initial phase are estimated at US$65.8 million and are illustrated in Table 21-4.

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Table 21-4:        On-siteInfrastructure Capital Cost Breakdown

WBS Description Initial Costs (US$M)
4100 Plant<br> Bulk Earthworks $10.6
4200 Plant<br> Power Switchyard and Power Distribution $10.2
4300 Fuel<br> Storage, Sewage, and Potable Water $6.6
4400 Water<br> Management 1.5
4500 Infrastructure<br> Buildings $5.0
4600 Tailings<br> Storage Facility $17.7
4700 Tailings<br> Pipeline $6.0
4800 Permanent<br> Camp $7.3
4900 Site<br> Services and Mobile Equipment $1.0
4000 On-Site Infrastructure $65.8

Note: Totals may not match due to rounding.

21.2.5.2 Off-site infrastructure

The off-site infrastructure costs for the initial phase consist of access road upgrades, water supply, power supply, and town relocation. The total off-site infrastructure costs are estimated at US$114.5 million and are illustrated in Table 21-5.

Table 21-5:        Off-siteInfrastructure Capital Cost Breakdown

WBS Description Initial Costs (US$M)
5100 Main<br> Access Road Upgrades $1.1
5200 Water<br> Supply $2.2
5300 Power<br> Supply (Transmission Line) $82.1
5400 Town<br> Relocation $29.2
5000 Off-Site Infrastructure $114.5

Note: Totals may not match due to rounding.

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21.2.6 Indirect Capital Costs
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Indirect costs include project preliminaries and project delivery (or EPCM). Project preliminaries include field indirects (temporary construction facilities, camp, and associated services), commissioning and operational readiness, vendor representatives, spares, and first fills. Project delivery includes engineering services and construction management. Project preliminaries for the initial phase are estimated at US$31.8 million while project delivery is estimated at US$45.6 million. Total indirect costs for the initial phase are estimated at US$77.4 million and are illustrated in Table 21-6.

Table 21-6:        IndirectCapital Cost Breakdown

WBS Description Initial Costs (US$M)
6100 Field<br> Indirects $17.1
6200 Commissioning<br> Operations Readiness $4.9
6300 Vendor<br> Representatives $3.7
6400 Spares $2.5
6500 First<br> Fills $3.7
6000 Project Preliminaries $31.8
7100 Engineering<br> Services $27.9
7200 Construction<br> Management $17.7
7000 Project Delivery **$45.**6
6000 + 7000 Indirect Costs **$77.**4

Note: Totals may not match due to rounding.

21.2.7 Owner (Corporate) Capital Costs

Owner costs for pre-production have been estimated by factors. The owner’s costs include:

· Owner’s project team and expenses
· Administration, finance, insurance and legal fees
--- ---
o Including pre-production general and administrative costs
--- ---
· Environmental consultation and management
--- ---
· Human resources, recruiting, and training
--- ---
· Permitting and regulatory compliance activities
--- ---
· Stakeholder relations
--- ---
· Site security.
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The owner’s costs for the initial phase are estimated at US$15.7 million.

21.2.8 Sustaining Capital
21.2.8.1 Mining Sustaining Costs
--- ---

The sustaining capital costs for mining are associated with site development and mine operations infrastructure. The total mining sustaining costs are estimated at US$14.6 million and are summarized in Table 21-7.

Table 21-7:        MineArea Capital Cost Summary

Item US$M
Site<br> Development Sustaining Capital Costs 13.5
Mine<br> Operations Infrastructure Sustaining Capital Costs 1.2
Total Sustaining Mining Capital 14.6

Note: Totals may not match due to rounding.

21.2.8.2 Infrastructure Sustaining Costs

The sustaining capital costs for infrastructure are associated with the expansion of TSF. The total sustaining costs are estimated at US$151.9 million and are summarized in Table 21-8.

Table 21-8:        InfrastructureSustaining Capital Costs

WBS Description Sustaining Costs (US$M)
4600 Tailings<br> Storage Facility $151.9
4000 Infrastructure Sustaining Costs **$**151.9

Note: Totals may not match due to rounding.

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21.2.9 Contingency
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Contingency costs account for the difference in costs between the estimated and actual cost of materials and equipment. The contingency is developed based upon the level of study and considers the level of project definition, the source or methodology of the estimates, and the expected accuracy range. It allows the capital estimate to include a provision to cover the risk from uncertainties that may arise in between the time the capital cost was developed compared to the actual costs during construction and pre-production.

The contingency for the Project has been built up by considering the cost inputs from Ausenco, MMTS, and NPM. The total contingency included in the initial capital costs is estimated at US$110.3 million or 23% of total direct and indirect costs, excluding capitalized mining opex.

21.2.10 Closure and Reclamation Planning

Closure and reclamation costs for the Project include allocations for:

· Process plant and on-site infrastructure
· TSF, WRSF, and Water Management
--- ---
· Mining, stockpiles, and haul roads
--- ---
· Long-term environmental monitoring
--- ---
· Long-term community considerations.
--- ---

The total closure and reclamation costs are estimated at US$149.8 million.

21.3 Operating Costs
21.3.1 Overview
--- ---

The total operating costs for the Project are estimated at US$16.93/t or US$4,258.4 million over the 19-year mine life. These operating costs do not include pre-production operating costs. A summary of operating costs is presented in Table 21-8.

Table 21-9:        OperatingCost Summary

Cost Area Total ($M) $/t % of Total
Mining 2,120.2 8.43 50
Process 1,791.2 7.12 42
G&A 347.0 1.38 8
Total 4,258.4 16.93 100
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21.3.2 Basis of Estimate
--- ---

The following was used to determine the project’s LOM process operating costs in agreement with the cost definition and estimate methodologies outlined below. This basis considers the development of a process plant designed to treat 22,000 t/d of mineralized material in the initial phase, and 44,000 t/d in the expansion phases. Process unit operations were benchmarked against similar or comparable processing plants to ensure accuracy of cost estimates.

Assumptions made in developing the process operating cost estimate are listed below:

· Mill production is designed to treat 22,000 t/d of mineralized material in the initial phase, and 44,000<br>t/d in the expansion phases.
· Process plant operating costs are calculated based on labor, power consumption, and process and maintenance<br>consumables.
--- ---
· Off-site gold refining, insurance, and transportation costs are excluded, as they are included elsewhere<br>in the financial model.
--- ---
· Labor rates were sourced from recent execution projects in the region.
--- ---
· Workforce will be comprised of local and regional workers.
--- ---
· Management and administrative staff will be on a 5/2 rotation (5 days in, 2 days out), whereas process<br>and maintenance staff will be on a 14/7 rotation (14 days in, 7 days out).
--- ---
· Management and administrative staff who are not required to be on site will be based out of La Pas or<br>will work remotely.
--- ---
· General and administration (G&A) costs were baselined against previous regional project experience.
--- ---
· No factor for spare parts has been applied to adjust for consumption of fewer spare parts in early years<br>of operation.
--- ---
· Grinding media consumption rates have been estimated based on the mill feed characteristics.
--- ---
· Reagent consumption rates have been estimated based on the metallurgical testwork results.
--- ---
· Reagents and consumable prices were obtained via recent regional quotations with location-specific freight<br>considerations.
--- ---
· Mobile equipment cost includes for fuel, maintenance, and lease price for the equipment.
--- ---
· The unit rate power cost of US$0.06/kWh from ENDE Transmisión via NPM.
--- ---
· The unit rate fuel cost of US$5.34/gallon (US$.1.41/L) based on current regional rates at the Project<br>site via NPM.
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21.3.3 Mine Operating Costs
--- ---

The mine operating cost estimates for the Carangas project are based on a mining contractor's quotation received in Q2 of 2026 and applied to the Carangas PEA mine production schedule.

The contractor provided the following unit rates for mining operations at Carangas, which have been directly applied to the scheduled material movement:

· Drilling and Blasting = $0.88/t
· Loading and Hauling = $1.58/t
--- ---
· Incremental Haul Distance per km, over 2 km = $0.17/t
--- ---
· Incremental Vertical Haul Distance per 10 m = $0.017/t
--- ---
· Rehandle Loading and Hauling = $1.58/t
--- ---

The contractor's unit rates incorporate a diesel price input of US$1.20/L.

Haulage profiles for all material sources to all destinations have been developed for each year of the mine production schedule. The annual average haul distances and vertical elevation changes have been measured and applied to the relevant incremental unit costs.

In addition, the operating cost estimate includes staff salaries and departmental overheads for an owner-employed team, covering roles such as project manager, contracts manager, technical services manager, geologists, surveyors, mine engineers, and a geotechnical engineer.

The estimated average unit mine operating costs are as follows:

· During the Pre-Production Construction Period = $2.89/t
o These costs are capitalized and included in the capital cost estimate for mining 19.1 Mt of pit rock.
--- ---
· From Years 1-16 = $3.11/t
--- ---
o This corresponds to a total cost of $2,058 million to move 663 Mt of rock
--- ---
· From Years 17-19 (stockpile rehandle period) = $1.62/t
--- ---
o This corresponds to a total cost of $62 million to rehandle 38 Mt of stockpiled low-grade and oxide material<br>to the mill.
--- ---

The total operating cost per tonne of mill feed is estimated at $8.43/t milled.

It is the QP’s opinion that the estimates are reasonable for the location and planned mine operation activities and can be utilized for a PEA.

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21.3.4 Process Operating Costs
--- ---

The process operation cost estimate includes costs relating to the operation of the mill. The process operating costs for the mill comprise of costs associated with reagent and consumable consumption, labor, process mobile equipment, power, and maintenance. The process operating costs for the initial phase are US$6.94/t and are illustrated in Table 21-9.

Table 21-10:      ProcessOperating Costs - Initial Phase - Concentrate Production

Cost Area Average Annual (US$M/a) US$/t milled % of Total
Power 13.2 1.65 23.7
Reagents 22.1 2.77 39.9
Consumables 10.8 1.35 19.5
Maintenance 4.1 0.51 7.4
Labour 3.2 0.39 5.7
Mobile<br> Equipment 1.6 0.20 2.9
Lab<br> Services 0.5 0.06 0.9
Water<br> Treatment - – -
Total 55.5 6.94 100.0

Note: Totals may not match due to rounding.

The process operating costs for the concentrate plant expansion phase are US$7.00/t and are illustrated in Table 21-10.

Table 21-11:      ProcessOperating Costs - Expansion Phase - Concentrate Production

Cost Area Average Annual (US$M/a) US$/t milled % of Total
Power 26.4 1.65 23.5
Reagents 44.3 2.77 39.5
Consumables 21.6 1.35 19.3
Maintenance 8.2 0.51 7.3
Labour 5.6 0.35 5.0
Mobile<br> Equipment 3.2 0.20 2.8
Lab<br> Services 0.9 0.06 0.8
Water<br> Treatment 1.9 0.12 1.7
Total 112.0 7.00 100.0

Note: Totals may not match due to rounding.

The process operating costs for the gold plant expansion phase are US$7.00/t and are illustrated in Table 21-11.

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Table 21-12:      ProcessOperating Costs - Expansion Phase – Doré Production

Cost Area Average Annual (US$M/a) US$/t milled % of Total
Power 12.3 1.54 20.7
Reagents 25.6 3.20 43.1
Consumables 10.9 1.36 18.3
Maintenance 4.5 0.57 7.6
Labour 2.5 0.31 4.2
Mobile<br> Equipment 1.5 0.19 2.6
Lab<br> Services 0.3 0.03 0.4
Water<br> Treatment 1.9 0.23 3.1
Total 59.5 7.43 100.0

Note: Totals may not match due to rounding.

21.3.5 General and Administrative Operating Costs

The G&A costs for the initial phase of the Project are estimated at US$1.67/t milled and are illustrated in Table 21-12.

Table 21-13:      Generaland Administrative Costs - Initial Phase

Cost Area Average Annual (US$M/a) US$/t milled % of Total
Site<br> Maintenance 2.1 0.26 15.7
Personnel 1.7 0.21 12.5
Human<br> Resources 1.2 0.15 9.2
Equipment 0.8 0.09 5.7
Health &<br> Safety 0.2 0.03 1.6
Environmental 0.6 0.07 4.2
IT &<br> Telecommunications 0.5 0.06 3.5
Contract<br> Services, Insurance, Legal 2.9 0.36 21.4
Camp<br> and Travel 2.8 0.35 20.8
Administrative<br> Costs 0.7 0.09 5.4
Total 13.3 1.67 100.0

Note: Totals may not match due to rounding.

The G&A costs for the concentrator expansion phase of the Project are estimated at US$1.24/t milled and are illustrated in Table 21-13.

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Table 21-14:      Generaland Administrative Costs - Concentrator Expansion Phase

Cost Area Average Annual (US$M/a) US$/t milled % of Total
Site<br> Maintenance 3.7 0.23 18.7
Personnel 2.2 0.14 11.2
Human<br> Resources 1.4 0.09 7.1
Equipment 1.5 0.10 7.8
Health &<br> Safety 0.3 0.02 1.5
Environmental 0.6 0.04 2.8
IT &<br> Telecommunications 0.7 0.04 3.6
Contract<br> Services, Insurance, Legal 4.1 0.25 20.4
Camp<br> and Travel 4.6 0.29 23.3
Administrative<br> Costs 0.7 0.05 3.7
Total 19.8 1.24 100.0

Note: Totals may not match due to rounding.

The G&A costs for the gold expansion phase of the Project are estimated at US$0.85/t milled and are illustrated in Table 21-14.

Table 21-15:      Generaland Administrative Costs - Gold Expansion Phase

Cost Area Average Annual (US$M/a) US$/t milled % of Total
Site<br> Maintenance 1.6 0.20 22.9
Personnel 0.5 0.07 8.0
Human<br> Resources 0.1 0.02 1.9
Equipment 0.8 0.10 11.5
Health &<br> Safety 0.2 0.02 2.7
Environmental 0.0 0.00 0.0
IT &<br> Telecommunications 0.3 0.03 3.8
Contract<br> Services, Insurance, Legal 1.2 0.15 17.6
Camp<br> and travel 1.9 0.24 27.8
Administrative<br> Costs 0.3 0.03 3.8
Total 6.8 0.85 100.0

Note: Totals may not match due to rounding.

21.3.6 Owner (Corporate) Operating Costs

The Owner’s costs for the Project are categorized as capital costs for the purpose of this study and are described in Section 21.2.7.

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21.3.7 Closure and Reclamation Planning
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The closure and reclamation costs for the Project are categorized as capital costs for the purpose of this study and are described in Section 21.2.10.

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22 Economic Analysis
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22.1 Forward-Looking Information
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The results of the economic analyses discussed in this section represent forward-looking information as defined under relevant securities law. The results depend on inputs that are subject to known and unknown risks, uncertainties, and other factors that may cause actual results to differ materially from those presented here. Information that is forward-looking includes the following:

· Mineral resource estimates
· Assumed commodity prices and exchange rates
--- ---
· Proposed mine production plan
--- ---
· Projected mining and process recovery rates
--- ---
· Assumptions as to mining dilution and estimated future production
--- ---
· Sustaining costs and proposed operating costs
--- ---
· Assumptions as to closure costs and closure requirements
--- ---
· Assumptions as to environmental, permitting, and social risks.
--- ---

Additional risks to the forward-looking information include:

· Changes to costs of production from what is assumed
· Unrecognized environmental risks
--- ---
· Unanticipated reclamation expenses
--- ---
· Unexpected variations in quantity of mineralized material, grade, or recovery rates
--- ---
· Accidents, labor disputes and other risks of the mining industry
--- ---
· Geotechnical or hydrogeological considerations during mining being different from what was assumed
--- ---
· Failure of mining methods to operate as anticipated
--- ---
· Failure of plant, equipment, or processes to operate as anticipated
--- ---
· Changes to assumptions as to the availability of electrical power, and the power rates used in the operating<br>cost estimates and financial analysis
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· Changes to site access, use of water for mining purposes and to time to obtain environment and other regulatory<br>permits
--- ---
· Ability to maintain the social license to operate
--- ---
· Changes to financing interest rates
--- ---
· Changes to applicable tax rates.
--- ---

The PEA is preliminary in nature, that it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the preliminary economic assessment will be realized

22.2 Methodologies Used

The Project has been evaluated using a discounted cashflow (DCF) analysis based on a 5% discount rate. Cash inflows consist of annual revenue projections. Cash outflows consist of capital expenditures, operating costs, taxes, and royalties. These are subtracted from the inflows to arrive at the annual cash flow projections. Cash flows are taken to occur at the midpoint of each period. It must be noted that tax calculations involve complex variables that can only be accurately determined during operations and, as such, the actual post-tax results may differ from those estimated. A sensitivity analysis was performed to assess the impact of variations in metal prices, discount rate, head grade, recovery, total operating cost, and total capital costs.

The capital and operating cost estimates developed specifically for this project are presented in Section 21 in Q2 2026 US dollars. The economic analysis has been run on a constant dollar basis with no inflation.

22.3 Financial Model Parameters
22.3.1 Assumptions
--- ---

The economic analysis was performed using a gold price of US$3,400/oz, a silver price of US$45.00/oz, a lead price of US$0.90/lb and a zinc price of US$1.20/lb. These metal prices were based on consensus analyst estimates and recently published economic studies. The forecasts used are meant to reflect the average metals price expectation over the life of the Project. No price inflation or escalation factors were considered. Commodity prices can be volatile, and there is the potential for deviation from the forecast.

The economic analysis also used the following assumptions:

· Construction and commissioning period of two years.
· Total mine life of 19 years.
--- ---
· Cost estimates in constant Q2 2026 US dollars with no inflation or escalation factors considered.
--- ---
· Results based on 100% ownership with a 6.0% NSR royalty applied to silver, and a 5.0% NSR royalty on gold,<br>lead, and zinc.
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· Capital costs funded with 100% equity.
--- ---
· All cash flows discounted to start of construction period using mid-period discounting convention.
--- ---
· All metal products are sold in the same year that they are produced.
--- ---
· Project revenue is derived from the sale of silver-lead concentrate, zinc concentrate, and gold doré.
--- ---
· No contractual arrangements for refining currently exist.
--- ---
22.3.2 Taxes
--- ---

The Project has been evaluated on an after-tax basis to provide an approximate value of the expected economics. The tax model was compiled by NPM. The tax model contains federal and state income tax computations based on Bolivia’s 2009 Constitution, 2014 Mining Law (Ley N° 535 de Minería y Metalurgia), general tax law (Law 843), and additional extractive taxes.

22.3.3 Working Capital

An estimate of working capital has been incorporated into the economic analysis based on the following assumptions.

Table 22-1:        WorkingCapital Assumptions

Description Units Value
Accounts<br> Receivable days 0
Inventory days 30
Accounts<br> Payable days 30
22.3.4 Closure Costs and Salvage Value
--- ---

Closure and salvage value are applied at the end of the LOM. Closure costs were estimated to be US$149.8 million with no salvage value assumed.

22.3.5 Royalties

Based on the agreements in place as of the date of this technical report and summarized in Section 4.5, an NSR royalties are applied to the recovered metals in the financial model. A base royalty of 5.0% is applied to the recovered lead and zinc, with 6.0% applied to the recovered silver, and 5.0% applied to the recovered gold.

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22.3.6 Off-site Costs
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The following off-site costs and sale terms are used for the economic analysis.

Table 22-2:        Off-TakeTerm Assumptions

Term Unit Value
Lead<br> Grade of Ag/Pb Concentrate % 24.0
Silver<br> Payability of Ag/Pb Concentrate % 96.5
Minimum<br> Silver Grade of Ag/Pb Concentrate g/t 50.0
Lead<br> Payability of Ag/Pb Concentrate % 95.0
Lead<br> Discount of Ag/Pb Concentrate % 3.0
Transport<br> and Marketing Cost of Ag/Pb Concentrate US$<br> /t 100.00
Treatment<br> Charge of Ag/Pb Concentrate US$<br> /t 120.00
Silver<br> Refining Charge of Ag/Pb Concentrate US$<br> /oz 0.50
Zinc<br> Grade of Ag/Zn Concentrate % 45.8
Silver<br> Payability of Ag/Zn Concentrate % 70.0
Minimum<br> Silver Grade of Ag/Zn Concentrate g/t 93.3
Zinc<br> Payability of Ag/Zn Concentrate % 85.0
Zinc<br> Discount of Ag/Zn Concentrate % 8.0
Transport<br> and Marketing Cost of Ag/Zn Concentrate US$<br> /t 175.00
Treatment<br> Charge of Ag/Zn Concentrate US$<br> /t 120.00
Silver<br> Refining Charge of Ag/Zn Concentrate US$<br> /oz 0.50
Silver<br> Recovery of Dore % 60.0
Gold<br> Recovery of Dore % 93.0
Silver<br> Payability of Dore % 95.0
Gold<br> Payability of Dore % 99.8
Silver<br> Refining, Transport, and Marketing Cost – Doré US$/oz 1.00
Gold<br> Refining, Transport, and Marketing Cost – Doré US$/oz 6.50
22.4 Economic Analysis
--- ---

The economic analysis is performed using a 5% discount rate. A 5% discount rate has been selected as 58% of gross revenue comes from silver, with 25% coming from gold, 11% coming from zinc, and 6% coming from lead. The pre-tax net present value discounted at 5% (NPV5%) is US$4,375 million, the IRR is 50.3%, and the payback period is 1.7 years from the start of commercial production. On a post-tax basis, the NPV5% is US$2,653 million, the IRR is 35.9%, and the payback period is 2.4 years from the start of commercial production. A summary of project economics is tabulated in Table 22-3.

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Table 22-3:        EconomicAnalysis Summary Table

Description Unit Value
General Inputs
Discount<br> rate % 5.0
Discount<br> convention Name Half<br> Period
Ag<br> price US$<br> /oz 45.00
Au<br> price US$<br> /oz 3,400.00
Pb<br> price US$<br> /lb 0.90
Zn<br> price US$<br> /lb 1.20
Exchange<br> rate US$:US$ 1.00
Construction<br> period Yr(s) 2
Working<br> capital - accounts payable Days –
Working<br> capital - accounts receivable Days 30
Working<br> capital - inventory Days 30
Average<br> Ag payability % 94.1%
Average<br> Au payability % 99.8%
Average<br> Pb payability % 84.8%
Average<br> Zn payability % 61.0%
Payable<br> AgEq koz 338,982
Payable<br> AuEq koz 4,487
Final Products
Ag-Pb<br> Concentrate Tonnage (dry) kt 2,032
Ag<br> recovered in Ag-Pb concentrate koz 184,950
Pb<br> recovered in Ag-Pb concentrate kt 488
Ag<br> payable in Ag-Pb concentrate koz 178,246
Pb<br> payable in Ag-Pb concentrate kt 427
Zn<br> Concentrate Tonnage (dry) kt 1,744
Ag<br> recovered in Zn concentrate koz 10,516
Zn<br> recovered in Zn concentrate kt 799
Ag<br> payable in Zn concentrate koz 5,635
Zn<br> payable in Zn concentrate kt 659
Ag<br> recovered in dore koz 11,832
Au<br> recovered in dore oz 1,144,282
Ag<br> payable in dore koz 11,241
Au<br> payable in dore oz 1,141,993
LOM Operating Costs
Mining<br> cost per tonne mined US$<br> /t 3.56
Mining<br> cost per tonne processed US$<br> /t 8.43
Processing<br> cost per tonne processed US$<br> /t 7.12
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Description Unit Value
--- --- ---
G&A<br> cost per tonne processed US$<br> /t 1.38
--- --- ---
Total<br> operating cost per tonne processed US$<br> /t 16.93
Refining &<br> transport cost US$<br> M 1,128.2
Base<br> Royalty US$<br> M 793.1
Total<br> cash cost (TCC)⁽¹⁾ US$/oz<br> Ag (1.51)
All-in<br> sustaining cost (AISC)⁽²⁾ US$/oz<br> Ag 0.11
Total<br> cash cost (TCC)⁽³⁾ US$/oz<br> AgEq 18.23
All-in<br> sustaining cost (AISC)⁽⁴⁾ US$/oz<br> AgEq 19.16
Total<br> cash cost (TCC)⁽³⁾ US$/oz<br> AuEq 1,377.38
All-in<br> sustaining cost (AISC)⁽⁴⁾ US$/oz<br> AuEq 1,447.87
Capital Costs
Initial<br> capital US$<br> M 644.5
Expansion<br> capital US$<br> M 422.7
Sustaining<br> capital US$<br> M 166.5
Closure<br> costs US$<br> M 149.8
Salvage<br> costs US$<br> M –
Capital<br> intensity US$/oz<br> Ag 3.30
Financials
Pre-tax<br> NPV (5%) US$<br> M 4,374.8
Pre-tax<br> IRR % 50.3
Pre-tax<br> payback Yr(s) 1.7
Post-tax<br> NPV (5%) US$<br> M 2,653.1
Post-tax<br> IRR % 35.9
Post-tax<br> payback Yr(s) 2.4
Revenue US$<br> M 15,254.2
EBITDA US$<br> M 9,074.5
EBITDA<br> margin % 59.5
Pre-tax<br> unlevered free cash flow US$<br> M 7,691.1
Post-tax<br> unlevered free cash flow US$<br> M 4,812.6
Taxes US$<br> M 2,878.6
Post-Tax<br> NPV (5%) / Initial Capital expenditure Ratio 4.1

Notes:

(1) TCC consist of mining costs, processing costs, mine-level G&A and refining charges and royalties less by-product credit

(2) AISC includes TCC plus sustaining capital, salvage value and closure costs

(3) TCC consist of mining costs, processing costs, mine-level G&A and refining charges and royalties

(4) AISC includes TCC plus sustaining capital, salvage value and closure costs

Gold equivalent ounces = gold mass + (silver mass * (silver price/gold price)) + (lead mass * (lead price/gold price)) + (zinc mass * (zinc price/gold price))

Silver equivalent ounces = silver mass + (gold mass * (gold price/silver price)) + (lead mass * (lead price/silver price)) + (zinc mass * (zinc price/silver price))

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Table 22-4:        CashflowStatement on an Annualized Basis

Units Total/Avg -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
Macro assumptions
Ag<br> price US$/oz 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45 45
Au<br> price US$/oz 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400 3,400
Pb<br> price US$/lb 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90 0.90
Zn<br> price US$/lb 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20 1.20
Free cash flow valuation
Revenue US$<br> M 15,254.2 – – 545.2 709.6 678.9 694.5 771.5 1,128.3 1,212.6 1,071.7 909.1 796.9 924.3 824.6 987.6 1,152.4 1,167.5 707.3 412.6 412.6 147.0 –
Operating<br> cost US$<br> M (4,258.4) – – (146.6) (144.3) (144.5) (143.5) (144.2) (380.2) (373.4) (355.1) (333.0) (322.2) (311.5) (305.0) (219.2) (198.4) (175.8) (161.6) (157.7) (157.7) (84.4) –
Refining<br> and transport charges US$<br> M (1,128.2) – – (52.7) (64.5) (55.7) (52.7) (66.4) (99.6) (92.1) (68.4) (47.0) (46.6) (44.1) (42.5) (40.7) (40.7) (40.0) (71.5) (83.5) (83.5) (36.0) –
Royalties US$<br> M (793.1) – – (28.8) (37.8) (36.6) (37.8) (41.4) (60.3) (66.0) (59.3) (49.4) (40.8) (47.3) (41.6) (50.4) (58.4) (58.8) (35.3) (18.5) (18.5) (6.1) –
EBITDA US$ M 9,074.5 – – 317.1 462.9 442.1 460.5 519.5 588.2 681.1 588.8 479.7 387.3 521.4 435.5 677.3 854.9 892.8 438.9 152.9 152.9 20.5 –
Initial<br> capital expenditure US$<br> M (644.5) (259.2) (385.2) – – – – – – – – – – – – – – – – – – – –
Expansionary<br> capital expenditure US$<br> M (422.7) – – – – – (176.3) (176.3) – – (35.0) (35.0) – – – – – – – – – – –
Sustaining<br> capital expenditure US$<br> M (166.5) – – (10.3) (9.0) (8.9) (8.9) (9.1) (11.2) (11.2) (10.7) (10.6) (10.1) (10.0) (10.0) (9.9) (7.7) (7.3) (7.3) (7.2) (7.2) (0.0) –
Closure<br> capital expenditure US$<br> M (149.8) – – – – – – – – – – – – – – – – – – – – (149.8) –
Salvage<br> value US$<br> M – – – – – – – – – – – – – – – – – – – – – – –
Changes<br> in working capital US$<br> M – – – – – – – – – – – – – – – – – – – – – – –
Pre-tax unlevered free cash flow US$ M 7,691.1 (259.2) (385.2) 306.8 453.9 433.2 275.3 334.1 577.0 669.9 543.2 434.0 377.3 511.4 425.5 667.4 847.3 885.5 431.6 145.7 145.7 (129.3) –
Pre-tax<br> cumulative unlevered free cash flow US$<br> M (259.2) (644.5) (337.6) 116.3 549.5 824.8 1,158.9 1,735.9 2,405.8 2,949.0 3,383.0 3,760.3 4,271.7 4,697.2 5,364.6 6,211.9 7,097.4 7,529.0 7,674.7 7,820.4 7,691.1 7,691.1
Unlevered<br> cash taxes US$<br> M (2,878.6) – – (89.3) (143.6) (135.4) (149.3) (171.1) (182.3) (216.7) (181.7) (155.2) (115.9) (166.4) (136.1) (216.5) (280.4) (292.8) (150.1) (48.1) (47.8) – –
Post-tax unlevered free cash flow US$ M 4,812.6 (259.2) (385.2) 217.6 310.3 297.7 126.0 163.0 394.7 453.3 361.4 278.9 261.3 345.0 289.4 451.0 566.9 592.7 281.6 97.6 97.9 (129.3) –
Post-tax<br> cumulative unlevered free cash flow US$<br> M (259.2) (644.5) (426.9) (116.6) 181.2 307.1 470.2 864.9 1,318.2 1,679.6 1,958.5 2,219.8 2,564.8 2,854.3 3,305.2 3,872.1 4,464.8 4,746.3 4,843.9 4,941.9 4,812.6 4,812.6
Production
Total<br> mineralized material mined kt 251,547 1,605 5,936 11,723 12,771 14,144 7,610 12,571 26,239 27,493 29,598 27,035 20,657 17,484 12,469 8,169 8,060 7,030 950 – – – –
Ag<br> head grade - total resource mined g/t 31.25 39.04 41.24 48.96 46.78 46.69 60.37 47.82 34.84 38.75 34.86 25.16 14.71 14.57 8.47 15.97 12.98 8.92 7.83 – – – –
Au<br> head grade - total resource mined g/t 0.20 0.01 0.02 0.01 0.02 0.02 0.01 0.01 0.04 0.07 0.13 0.14 0.21 0.32 0.39 0.71 0.95 1.16 1.86 – – – –
Pb<br> head grade - total resource mined % 0.31% 0.62% 0.50% 0.47% 0.47% 0.37% 0.36% 0.40% 0.31% 0.29% 0.24% 0.31% 0.36% 0.27% 0.24% 0.10% 0.08% 0.07% 0.03% – – – –
Zn<br> head grade - total resource mined % 0.57% 0.05% 0.13% 0.75% 0.70% 0.61% 0.79% 0.84% 0.72% 0.60% 0.53% 0.61% 0.64% 0.51% 0.51% 0.21% 0.16% 0.12% 0.04% – – – –
Contained<br> metal - Ag koz 252,724 2,015 7,871 18,454 19,208 21,233 14,771 19,329 29,389 34,247 33,175 21,869 9,767 8,187 3,396 4,194 3,365 2,015 239 – – – –
Contained<br> metal - Au oz 1,597,120 341 3,741 2,767 6,865 7,174 2,903 5,669 31,598 58,692 126,844 123,146 138,121 179,625 156,524 187,766 246,458 261,967 56,921 – – – –
Contained<br> metal - Pb kt 772 10 29 55 60 52 27 51 81 79 72 85 75 48 30 8 6 5 0 – – – –
Contained<br> metal - Zn kt 1,438 1 8 88 90 86 60 106 188 165 158 165 132 90 64 17 13 9 0 – – – –
Total<br> waste mined kt 363,429 3,395 8,181 16,840 13,386 12,355 16,636 12,428 58,761 50,890 38,258 32,383 30,343 27,516 27,610 9,288 4,149 1,006 3 – – – –
Total<br> material mined kt 614,976 5,000 14,117 28,563 26,157 26,500 24,247 25,000 85,000 78,383 67,856 59,418 51,000 45,000 40,080 17,457 12,209 8,036 953 – – – –
Total<br> material mined (excl. pre-strip) kt 595,859 – – 28,563 26,157 26,500 24,247 25,000 85,000 78,383 67,856 59,418 51,000 45,000 40,080 17,457 12,209 8,036 953 – – – –
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Units Total/Avg -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
--- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- ---
Processed
Mill<br> feed kt 251,548 – – 7,250 8,000 8,000 8,000 8,000 16,000 16,000 16,000 15,000 16,000 16,000 16,000 16,000 16,000 15,000 16,000 16,000 16,000 6,298 –
Mill<br> head grade – Ag g/t 31.25 – – 49.77 58.80 58.69 61.49 64.51 48.45 54.18 51.46 38.45 19.52 20.06 15.80 19.47 17.84 16.28 20.41 12.52 12.52 8.80 –
Mill<br> head grade – Au g/t 0.15 – – – – – – – – – – 0.12 0.22 0.31 0.29 0.36 0.48 0.54 0.11 – – – –
Mill<br> head grade – Pb % 0.27% – – 0.50% 0.51% 0.42% 0.37% 0.49% 0.36% 0.35% 0.27% 0.20% 0.18% 0.16% 0.15% 0.14% 0.14% 0.15% 0.27% 0.33% 0.33% 0.31% –
Mill<br> head grade – Zn % 0.51% – – 0.87% 0.91% 0.79% 0.73% 0.94% 0.71% 0.59% 0.43% 0.30% 0.31% 0.30% 0.32% 0.30% 0.30% 0.32% 0.56% 0.65% 0.65% 0.70% –
Contained<br> – Ag koz 252,725 – – 11,602 15,123 15,095 15,814 16,593 24,923 27,870 26,470 18,544 10,043 10,317 8,128 10,016 9,180 7,850 10,499 6,439 6,439 1,782 –
Contained<br> – Au oz 1,230,411 – – – – – – – – – – 56,840 115,206 156,896 148,938 187,322 246,368 261,920 56,921 – – – –
Contained<br> – Pb kt 682 – – 36 40 34 30 39 58 57 43 30 28 26 24 22 23 22 44 53 53 20 –
Contained<br> – Zn kt 1,286 – – 63 72 63 58 75 114 95 68 45 50 48 51 47 47 47 90 104 104 44 –
Total<br> Ag contained in Silver Zone koz 233,005 – – 11,602 15,123 15,095 15,814 16,593 24,923 27,870 26,470 17,071 7,281 7,101 5,624 5,853 5,827 5,839 10,260 6,439 6,439 1,782 –
Total<br> Au contained in Gold Zone oz 1,230,411 – – – – – – – – – – 56,840 115,206 156,896 148,938 187,322 246,368 261,920 56,921 – – – –
Ag-Pb Concentrate
Tonnage<br> – dry kt 2,032 – – 104 121 100 90 119 173 172 127 88 85 78 71 65 66 65 127 158 158 66 –
Recovered<br> – Ag koz 184,950 – – 9,589 12,570 12,276 12,712 13,763 19,945 22,275 20,422 13,119 5,815 5,611 4,396 4,531 4,512 4,521 7,972 4,793 4,793 1,336 –
Recovered<br> – Pb kt 488 – – 25 29 24 22 28 42 41 31 21 20 19 17 16 16 16 30 38 38 16 –
Recovered<br> – Zn kt 282 – – 14 16 14 13 16 25 21 15 10 11 11 11 10 10 10 20 23 23 10 –
Payable<br> – Ag koz 178,246 – – 9,254 12,130 11,847 12,267 13,281 19,247 21,495 19,707 12,659 5,611 5,415 4,242 4,372 4,354 4,363 7,692 4,539 4,539 1,231 –
Payable<br> – Pb kt 427 – – 22 25 21 19 25 36 36 27 19 18 16 15 14 14 14 27 33 33 14 –
Payable<br> – Zn kt – – – – – – – – – – – – – – – – – – – – – – –
Ag-Zn Concentrate
Tonnage<br> – dry kt 1,744 – – 76 96 84 81 102 158 131 91 60 69 65 69 64 64 64 121 143 143 64 –
Recovered<br> – Ag koz 10,516 – – 439 666 695 775 802 1,187 1,360 1,290 821 322 310 228 239 238 239 403 204 204 92 –
Recovered<br> – Pb kt 16 – – 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 0 –
Recovered<br> – Zn kt 799 – – 35 44 39 37 47 72 60 42 27 32 30 32 29 29 29 55 65 65 29 –
Payable<br> – Ag koz 5,635 – – 211 377 442 531 496 714 952 903 575 114 114 21 49 47 48 41 – – – –
Payable<br> – Pb kt – – – – – – – – – – – – – – – – – – – – – – –
Payable<br> – Zn kt 659 – – 29 36 32 31 39 60 50 34 23 26 25 26 24 24 24 46 54 54 24 –
Dore
Recovered<br> – Ag koz 11,832 – – – – – – – – – – 884 1,657 1,929 1,502 2,498 2,012 1,206 143 – – – –
Recovered<br> – Au koz 1,144,282 – – – – – – – – – – 52,861 107,142 145,914 138,512 174,209 229,122 243,586 52,937 – – – –
Payable<br> – Ag koz 11,241 – – – – – – – – – – 839 1,574 1,833 1,427 2,373 1,911 1,146 136 – – – –
Payable<br> – Au koz 1,141,993 – – – – – – – – – – 52,755 106,927 145,622 138,235 173,861 228,664 243,099 52,831 – – – –
Revenue
Total payable – Ag koz 195,121 – – 9,465 12,506 12,289 12,798 13,777 19,961 22,448 20,610 14,074 7,300 7,362 5,690 6,794 6,313 5,557 7,870 4,539 4,539 1,231 –
Total payable – Au oz 1,141,993 – – – – – – – – – – 52,755 106,927 145,622 138,235 173,861 228,664 243,099 52,831 – – – –
Total payable – Pb kt 427 – – 22 25 21 19 25 36 36 27 19 18 16 15 14 14 14 27 33 33 14 –
Total payable – Zn kt 659 – – 29 36 32 31 39 60 50 34 23 26 25 26 24 24 24 46 54 54 24 –
Total payable – AgEq koz 338,982 – – 12,116 15,769 15,088 15,433 17,145 25,073 26,946 23,815 20,202 17,708 20,540 18,325 21,947 25,609 25,944 15,719 9,168 9,168 3,267 –
Total payable – AuEq koz 4,487 – – 160 209 200 204 227 332 357 315 267 234 272 243 290 339 343 208 121 121 43 –
Ag gross revenue US$ M 8,780.5 – – 426 563 553 576 620 898 1,010 927 633 328 331 256 306 284 250 354 204 204 55 –
Au gross revenue US$ M 3,882.8 – – – – – – – – – – 179 364 495 470 591 777 827 180 – – – –
Pb gross revenue US$ M 846.8 – – 43 50 42 38 49 72 71 53 37 36 32 30 27 27 27 53 66 66 27 –
Zn gross revenue US$ M 1,744.1 – – 76 96 84 81 102 158 131 91 60 69 65 69 64 64 64 121 143 143 64 –
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Units Total/Avg -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20
--- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- --- ---
Operating costs
Mine<br> operating costs US$<br> M 2,120.2 – – 82.1 75.5 75.7 74.7 75.4 248.4 241.6 223.3 205.2 187.1 176.4 169.9 84.1 63.3 46.8 28.9 25.9 25.9 10.4 –
Process<br> plant operating costs US$<br> M 1,791.2 – – 51.2 55.5 55.5 55.5 55.5 112.0 112.0 112.0 107.7 115.0 115.0 115.0 115.0 115.0 108.9 112.6 112.0 112.0 54.2 –
General<br> and Administration US$<br> M 347.0 – – 13.3 13.3 13.3 13.3 13.3 19.8 19.8 19.8 20.2 20.2 20.2 20.2 20.2 20.2 20.2 20.2 19.8 19.8 19.8 –
Total<br> OPEX US$<br> M 4,258.4 – – 146.6 144.3 144.5 143.5 144.2 380.2 373.4 355.1 333.0 322.2 311.5 305.0 219.2 198.4 175.8 161.6 157.7 157.7 84.4 –
Off-site & royalties
Off-site<br> charges US$<br> M 1,128.2 – – 52.7 64.5 55.7 52.7 66.4 99.6 92.1 68.4 47.0 46.6 44.1 42.5 40.7 40.7 40.0 71.5 83.5 83.5 36.0 –
Royalties US$<br> M 793.1 – – 28.8 37.8 36.6 37.8 41.4 60.3 66.0 59.3 49.4 40.8 47.3 41.6 50.4 58.4 58.8 35.3 18.5 18.5 6.1 –
Cash costs
Total<br> cash cost (TCC)⁽¹⁾ US$/oz<br> Ag (1.51) – – 11.50 7.98 9.03 9.02 7.29 15.53 14.66 16.43 10.92 (8.06) (25.83) (31.54) (54.69) (90.43) (115.67) (10.77) 11.31 11.31 28.33 –
All-in<br> sustaining cost (AISC)⁽²⁾ US$/oz<br> Ag 0.11 – – 12.58 8.70 9.75 9.71 7.95 16.09 15.16 16.95 11.67 (6.68) (24.47) (29.78) (53.24) (89.21) (114.35) (9.85) 12.90 12.90 150.05 –
Capital expenditures
Initial capital US$ M 644.5 259.2 385.2 – – – – – – – – – – – – – – – – – – – –
Mining US$<br> M 58.7 16.9 41.8 – – – – – – – – – – – – – – – – – – – –
Concentrate<br> plant US$<br> M 202.1 70.7 131.4 – – – – – – – – – – – – – – – – – – – –
Gold<br> plant US$<br> M – – – – – – – – – – – – – – – – – – – – – – –
On-site<br> infrastructure US$<br> M 65.8 19.7 46.1 – – – – – – – – – – – – – – – – – – – –
Off-site<br> infrastructure US$<br> M 114.5 34.3 80.1 – – – – – – – – – – – – – – – – – – – –
Project<br> preliminaries US$<br> M 31.7 9.5 22.2 – – – – – – – – – – – – – – – – – – – –
Project<br> delivery US$<br> M 45.6 13.7 31.9 – – – – – – – – – – – – – – – – – – – –
Owner's<br> costs US$<br> M 15.7 4.7 11.0 – – – – – – – – – – – – – – – – – – – –
Contingency US$<br> M 110.3 89.5 20.7 – – – – – – – – – – – – – – – – – – – –
Expansion capital US$ M 422.7 – – – – – 176.3 176.3 – – 35.0 35.0 – – – – – – – – – – –
Mining US$<br> M – – – – – – – – – – – – – – – – – – – – – – –
Concentrate<br> plant US$<br> M 202.1 – – – – – 101.1 101.1 – – – – – – – – – – – – – – –
Gold<br> plant US$<br> M 42.9 – – – – – – – – – 21.5 21.5 – – – – – – – – – – –
On-site<br> infrastructure US$<br> M 21.1 – – – – – 10.2 10.2 – – 0.4 0.4 – – – – – – – – – – –
Off-site<br> infrastructure US$<br> M - – – – – – - - – – – – – – – – – – – – – – –
Project<br> preliminaries US$<br> M 20.0 – – – – – 8.3 8.3 – – 1.6 1.6 – – – – – – – – – – –
Project<br> delivery US$<br> M 46.8 – – – – – 19.4 19.4 – – 3.9 3.9 – – – – – – – – – – –
Owner's<br> costs US$<br> M – – – – – – – – – – – – – – – – – – – – – – –
Contingency US$<br> M 89.7 – – – – – 37.3 37.3 – – 7.6 7.6 – – – – – – – – – – –
Sustaining capital US$ M 166.5 – – 10.3 9.0 8.9 8.9 9.1 11.2 11.2 10.7 10.6 10.1 10.0 10.0 9.9 7.7 7.3 7.3 7.2 7.2 0.0 –
Mining US$<br> M 14.6 – – 1.4 1.3 1.2 1.2 1.4 1.3 1.3 0.8 0.8 0.7 0.7 0.7 0.6 0.6 0.2 0.1 0.1 0.1 0.0 –
On-site<br> infrastructure US$<br> M 151.9 – – 8.9 7.7 7.7 7.7 7.7 9.8 9.8 9.8 9.8 9.3 9.3 9.3 9.3 7.1 7.1 7.1 7.1 7.1 – –
Closure<br> cost US$<br> M 149.8 – – – – – – – – – – – – – – – – – – – – 149.8 –
Total capital expenditures US$ M 1,383.4 259.2 385.2 10.3 9.0 8.9 185.2 185.4 11.2 11.2 45.7 45.7 10.1 10.0 10.0 9.9 7.7 7.3 7.3 7.2 7.2 149.8 –

Notes:

(1) TCC consist of mining costs, processing costs, mine-level G&A and refining charges and royalties less by-product credit.

(2) AISC includes TCC plus sustaining capital, salvage value and closure costs.

Gold equivalent ounces = gold mass + (silver mass * (silver price/gold price)) + (lead mass * (lead price/gold price)) + (zinc mass * (zinc price/gold price))

Silver equivalent ounces = silver mass + (gold mass * (gold price/silver price)) + (lead mass * (lead price/silver price)) + (zinc mass * (zinc price/silver price))

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22.5 Sensitivity Analysis
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A sensitivity analysis was conducted on the base-case pre-tax and post-tax NPV5% and IRR of the project using the following variables: metal prices, discount rate, total operating costs, initial capital costs, recovery, and head grade.

As shown in Figure 22-1, the sensitivity analysis reveals that the project is most sensitive to changes in recovery, head grade and commodity prices, and less sensitive to operating costs, and initial capital costs.

Table 22-5 shows the pre-tax sensitivity analysis results while Table 22-6 shows post-tax sensitivity analysis results.

Figure 22-1:      SensitivityAnalysis

Source: Ausenco, 2026

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Table 22-5:        Pre-TaxSensitivity

Pre-Tax<br> Sensitivity to Metal Prices
Pre-Tax<br> NPV Sensitivity to Discount Rate Pre-Tax<br> IRR Sensitivity to Discount Rate
Commodity Price Commodity Price
4,378 (20.0%) (10.0%) – 10.0% 20.0% Discount Rate 50.2% (20.0%) (10.0%) – 10.0% 20.0%
3.0% 3,347.1 4,395.7 5,444.3 6,492.9 7,541.6 3.0% 35.9% 43.4% 50.3% 56.7% 62.9%
5.0% 2,651.0 3,512.9 4,374.8 5,236.8 6,098.7 5.0% 35.9% 43.4% 50.3% 56.7% 62.9%
8.0% 1,887.9 2,544.2 3,200.6 3,856.9 4,513.2 8.0% 35.9% 43.4% 50.3% 56.7% 62.9%
10.0% 1,513.3 2,067.9 2,622.5 3,177.1 3,731.6 10.0% 35.9% 43.4% 50.3% 56.7% 62.9%
12.0% 1,216.1 1,689.3 2,162.5 2,635.7 3,108.9 12.0% 35.9% 43.4% 50.3% 56.7% 62.9%
Pre-Tax<br> NPV Sensitivity to Opex Pre-Tax<br> IRR Sensitivity to Opex
Commodity Price Commodity Price
4,378 (20.0%) (10.0%) – 10.0% 20.0% Opex 50.2% (20.0%) (10.0%) – 10.0% 20.0%
(20.0%) 3,156.9 4,018.9 4,880.8 5,742.8 6,604.7 (20.0%) 40.0% 47.0% 53.7% 60.0% 66.0%
(10.0%) 2,903.9 3,765.9 4,627.8 5,489.8 6,351.7 (10.0%) 38.0% 45.2% 52.0% 58.4% 64.4%
– 2,651.0 3,512.9 4,374.8 5,236.8 6,098.7 – 35.9% 43.4% 50.3% 56.7% 62.9%
10.0% 2,398.0 3,259.9 4,121.9 4,983.8 5,845.7 10.0% 33.8% 41.4% 48.5% 55.1% 61.4%
20.0% 2,145.0 3,006.9 3,868.9 4,730.8 5,592.8 20.0% 31.6% 39.5% 46.7% 53.4% 59.8%
Pre-Tax<br> NPV Sensitivity to Initial Capex Pre-Tax<br> IRR Sensitivity to Initial Capex
Commodity Price Commodity Price
4,378 (20.0%) (10.0%) – 10.0% 20.0% Initial Capex 50.2% (20.0%) (10.0%) – 10.0% 20.0%
(20.0%) 2,773.1 3,635.1 4,497.0 5,359.0 6,220.9 (20.0%) 23.8% 30.8% 37.2% 43.1% 48.7%
(10.0%) 2,712.1 3,574.0 4,435.9 5,297.9 6,159.8 (10.0%) 30.1% 37.3% 44.0% 50.2% 56.0%
– 2,651.0 3,512.9 4,374.8 5,236.8 6,098.7 – 35.9% 43.4% 50.3% 56.7% 62.9%
10.0% 2,589.9 3,451.8 4,313.7 5,175.7 6,037.6 10.0% 41.3% 49.0% 56.2% 63.0% 69.4%
20.0% 2,528.8 3,390.7 4,252.6 5,114.6 5,976.5 20.0% 46.4% 54.4% 61.8% 68.9% 75.7%
Pre-Tax<br> NPV Sensitivity to Mill Head Grade Pre-Tax<br> IRR Sensitivity to Mill Head Grade
Commodity Price Commodity Price
4,378 (20.0%) (10.0%) – 10.0% 20.0% Mill Head Grade 50.2% (20.0%) (10.0%) – 10.0% 20.0%
(20.0%) 1,399.9 2,089.4 2,779.0 3,468.6 4,158.1 (20.0%) 23.8% 30.8% 37.2% 43.1% 48.7%
(10.0%) 2,025.4 2,801.2 3,576.9 4,352.7 5,128.4 (10.0%) 30.1% 37.3% 44.0% 50.2% 56.0%
– 2,651.0 3,512.9 4,374.8 5,236.8 6,098.7 – 35.9% 43.4% 50.3% 56.7% 62.9%
10.0% 3,276.5 4,224.6 5,172.8 6,120.9 7,069.0 10.0% 41.3% 49.0% 56.2% 63.0% 69.4%
20.0% 3,902.0 4,936.3 5,970.7 7,005.0 8,039.3 20.0% 46.4% 54.4% 61.8% 68.9% 75.7%
Pre-Tax<br> NPV Sensitivity to Recovery Pre-Tax<br> IRR Sensitivity to Recovery
Commodity Price Commodity Price
4,378 (20.0%) (10.0%) – 10.0% 20.0% Recovery 50.2% (20.0%) (10.0%) – 10.0% 20.0%
(20.0%) 1,399.9 2,089.4 2,779.0 3,468.6 4,158.1 (20.0%) 23.8% 30.8% 37.2% 43.1% 48.7%
(10.0%) 2,025.4 2,801.2 3,576.9 4,352.7 5,128.4 (10.0%) 30.1% 37.3% 44.0% 50.2% 56.0%
– 2,651.0 3,512.9 4,374.8 5,236.8 6,098.7 – 35.9% 43.4% 50.3% 56.7% 62.9%
10.0% 3,240.3 4,183.9 5,127.5 6,071.1 7,014.8 10.0% 41.3% 49.0% 56.2% 63.0% 69.4%
20.0% 3,719.6 4,731.1 5,742.6 6,754.1 7,765.6 20.0% 46.2% 54.2% 61.8% 68.9% 75.6%
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Table 22-6:        Post-TaxSensitivity

Post-Tax<br> Sensitivity to Metal Prices
Post-Tax<br> NPV Sensitivity to Discount Rate Post-Tax<br> IRR Sensitivity to Discount Rate
Commodity Price Commodity Price
2,655 (20.0%) (10.0%) – 10.0% 20.0% Discount Rate 35.9% (20.0%) (10.0%) – 10.0% 20.0%
3.0% 2,004.9 2,677.1 3,349.1 4,013.8 4,609.4 3.0% 25.5% 30.9% 35.9% 40.6% 44.6%
5.0% 1,550.7 2,101.9 2,653.1 3,198.3 3,689.3 5.0% 25.5% 30.9% 35.9% 40.6% 44.6%
8.0% 1,053.3 1,471.7 1,890.0 2,304.0 2,679.1 8.0% 25.5% 30.9% 35.9% 40.6% 44.6%
10.0% 809.6 1,162.5 1,515.3 1,864.4 2,181.9 10.0% 25.5% 30.9% 35.9% 40.6% 44.6%
12.0% 616.8 917.4 1,217.9 1,515.3 1,786.3 12.0% 25.5% 30.9% 35.9% 40.6% 44.6%
Post-Tax<br> NPV Sensitivity to Opex Post-Tax<br> IRR Sensitivity to Opex
Commodity Price Commodity Price
2,655 (20.0%) (10.0%) – 10.0% 20.0% Opex 35.9% (20.0%) (10.0%) – 10.0% 20.0%
(20.0%) 1,875.4 2,426.5 2,956.9 3,432.5 3,879.1 (20.0%) 28.5% 33.6% 38.4% 42.5% 46.2%
(10.0%) 1,713.0 2,264.2 2,815.4 3,323.3 3,792.0 (10.0%) 27.0% 32.3% 37.2% 41.6% 45.4%
– 1,550.7 2,101.9 2,653.1 3,198.3 3,689.3 – 25.4% 30.9% 35.9% 40.6% 44.6%
10.0% 1,388.3 1,939.6 2,490.8 3,041.9 3,568.9 10.0% 23.8% 29.4% 34.6% 39.4% 43.7%
20.0% 1,226.0 1,777.3 2,328.5 2,879.6 3,426.8 20.0% 22.2% 27.9% 33.2% 38.1% 42.7%
Post-Tax<br> NPV Sensitivity to Initial Capex Post-Tax<br> IRR Sensitivity to Initial Capex
Commodity Price Commodity Price
2,655 (20.0%) (10.0%) – 10.0% 20.0% Initial Capex 35.9% (20.0%) (10.0%) – 10.0% 20.0%
(20.0%) 1,672.9 2,224.1 2,775.3 3,303.7 3,791.2 (20.0%) 31.0% 37.3% 43.1% 48.1% 52.8%
(10.0%) 1,611.8 2,163.0 2,714.2 3,251.0 3,740.3 (10.0%) 27.9% 33.8% 39.1% 44.0% 48.3%
– 1,550.7 2,101.9 2,653.1 3,198.3 3,689.3 – 25.4% 30.9% 35.9% 40.6% 44.6%
10.0% 1,489.6 2,040.8 2,592.0 3,141.7 3,637.4 10.0% 23.3% 28.4% 33.2% 37.6% 41.5%
20.0% 1,428.5 1,979.7 2,530.9 3,082.0 3,585.4 20.0% 21.5% 26.4% 30.8% 35.0% 38.8%
Post-Tax<br> NPV Sensitivity to Mill Head Grade Post-Tax<br> IRR Sensitivity to Mill Head Grade
Commodity Price Commodity Price
2,655 (20.0%) (10.0%) – 10.0% 20.0% Mill Head Grade 35.9% (20.0%) (10.0%) – 10.0% 20.0%
(20.0%) 750.2 1,191.2 1,632.2 2,073.1 2,514.0 (20.0%) 16.5% 21.7% 26.4% 30.7% 34.8%
(10.0%) 1,150.4 1,646.5 2,142.6 2,638.7 3,130.6 (10.0%) 21.2% 26.5% 31.3% 35.8% 40.1%
– 1,550.7 2,101.9 2,653.1 3,198.3 3,689.3 – 25.5% 30.9% 35.9% 40.6% 44.6%
10.0% 1,950.9 2,557.3 3,163.5 3,707.8 4,216.3 10.0% 29.4% 35.0% 40.3% 44.7% 48.8%
20.0% 2,351.2 3,012.6 3,629.7 4,192.9 4,725.2 20.0% 33.1% 38.9% 44.0% 48.6% 52.7%
Post-Tax<br> NPV Sensitivity to Recovery Post-Tax<br> IRR Sensitivity to Recovery
Commodity Price Commodity Price
2,655 (20.0%) (10.0%) – 10.0% 20.0% Recovery 35.9% (20.0%) (10.0%) – 10.0% 20.0%
(20.0%) 750.2 1,191.2 1,632.2 2,073.1 2,514.0 (20.0%) 16.5% 21.7% 26.4% 30.7% 34.8%
(10.0%) 1,150.4 1,646.5 2,142.6 2,638.7 3,130.6 (10.0%) 21.2% 26.5% 31.3% 35.8% 40.1%
– 1,550.7 2,101.9 2,653.1 3,198.3 3,689.3 – 25.5% 30.9% 35.9% 40.6% 44.6%
10.0% 1,926.8 2,530.0 3,133.3 3,679.6 4,187.3 10.0% 29.3% 35.0% 40.2% 44.7% 48.8%
20.0% 2,229.3 2,875.3 3,497.0 4,055.4 4,580.2 20.0% 32.9% 38.7% 43.9% 48.4% 52.6%
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23 Adjacent Properties
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This section is not relevant to this report. There are no properties directly adjacent to Carangas.

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24 Other Relevant Data and Information
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No other relevant data or information has been included in this Report.

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25 Interpretation and Conclusions
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25.1 Introduction
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The QPs note the following interpretations and conclusions in their respective areas of expertise, based on the review of data available for this Report.

25.2 Mineral Tenure, Surface Rights, Water Rights, Royalties and Agreements

Based on the information reviewed for this report, the Carangas Project is considered to have secure mineral exploration rights through its valid Prospecting and Exploration Licenses (PELs), and all statutory obligations appear to be current. However, because mineral rights in Bolivia do not include ownership of the land surface, continued access to the Property depends on maintaining agreements with the local communities who own the surface rights. NPM has successfully obtained the necessary permissions for exploration, including drilling and the construction of temporary infrastructure, and no issues have been identified that would affect the validity of the current exploration work. Environmental baseline work and permitting requirements for future stages are well understood; however, any move toward development will require additional approvals and updated land use agreements. Overall, the Property tenure is secure for ongoing exploration; however, the Project’s ability to advance will continue to rely on maintaining strong relationships with the surrounding communities and meeting all regulatory, environmental, and social obligations as work progresses.

25.3 Geology and Mineralization

Carangas is a large silver-gold-lead-zinc polymetallic deposit hosted in a caldera-diatreme volcanic complex of the Tertiary age in the South American Epithermal-Porphyry Belt. Controlled by the temperature and pressure of the underlying hydrothermal system, mineralization is zoned into separate zones: a near-surface Upper Silver Zone dominated by silver plus a moderate amount of lead and zinc, a Middle Zinc Zone dominated by zinc plus a minor amount of silver and lead, and a Lower Gold Zone dominated by gold plus small amounts of silver, copper and zinc. Gold mineralization remains open to the north and northeast directions at depth. Beyond the drilled area, multiple IP chargeability anomalies exhibit geophysical signatures similar to those of the known mineralization. These anomalies constitute targets for future drilling to assess if additional material is suitable for consideration in Mineral Resources.

25.4 Exploration

Previous work at the Carangas Project demonstrated the potential for expansion of the Mineral Resources. Gold mineralization remains open to the north and northeast directions and at depth. Below the conceptual pit constraint, gold-dominated mineralized material of similar size and grade to the reported Mineral Resources of the Gold Domain exists within the conceptual pit.

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Beyond the drilled area, multiple IP chargeability anomalies with geophysical signatures similar to those of the known mineralization exist. These anomalies constitute targets for future drilling to assess whether additional material suitable for consideration in Mineral Resources exists.

The QP notes that at this stage, any of the IP chargeability anomaly prospects mentioned in this report have not been reported as Mineral Resources, and there is no guarantee that, through further exploration, Mineral Resources will be defined.

25.5 Drilling

Drilling completed at the Carangas Project has successfully outlined a broad epithermal system characterized by near-surface silver mineralization and a deeper, gold-dominant zone. The drill programs have provided consistent down-hole intercepts that confirm the presence of extensive mineralized bodies across the Central Valley area and around the volcanic domes. Overall, the drilling has reliably defined the lateral continuity of mineralization within the current drilling pattern, supporting the geological model used for the Project’s Mineral Resource estimates.

Despite this positive outcome, the orientation and structural controls of the mineralized zones remain only partially understood. The majority of holes were drilled without oriented core, and the geological model, while robust at a deposit scale, does not yet fully constrain the true geometry of the mineralization. As a result, the relationship between the reported sample lengths and the true thickness of the mineralized intervals cannot be definitively established. All drill intersections therefore represent down-hole lengths, and true thickness is expected to vary depending on local lithology, breccia development, and the configuration of the hydrothermal system.

The available drilling is adequate to support the current Mineral Resource classification and the conclusions of the PEA, however, additional data will be required to refine the understanding of mineralized controls. Further drilling with systematic structural logging, including the collection of oriented core and closer-spaced infill drilling, will be essential to improve confidence in the geometry of the mineralized zones. These steps will improve the resolution of true thicknesses, enhance domain definitions, and provide the level of geological support required for future upgrades to the Mineral Resource categories and for potential feasibility-level studies.

25.6 Sample Preparation, Analyses and Security

Based on the review of sampling, preparation, analytical, and security procedures applied at the Carangas Project, the data collection process is considered reasonable and consistent with industry best practices. Drill core sampling, handling, and preparation were carried out under controlled conditions by trained personnel, with clear procedures in place for logging, cutting, bagging, transport, and storage. The chain of custody was well maintained from drill site to laboratory, and no material issues related to sample security or integrity were identified.

The analytical program incorporated a comprehensive QA/QC framework that follows typical industry benchmarks, including the regular insertion of certified reference materials, blanks, and duplicates. The performance of these control samples demonstrates good analytical accuracy and precision for the key elements of interest, particularly silver and gold. Although some variability was observed in gold assays, likely related to the presence of coarse or free gold, this behavior is typical for epithermal systems and was appropriately monitored and managed through the QA/QC program.

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Overall, the QP considers the sample preparation, analytical methods, QA/QC procedures, and security protocols to be appropriate and reliable for use in Mineral Resource estimation.

25.7 Metallurgical Testwork

The metallurgical testwork programs in Section 13 were carried out by Bureau Veritas Minerals in Richmond, British Columbia, Canada and ALS Metallurgy in Kamloops, British Columbia, Canada under the supervision of Dr. Jinxing Ji on behalf of NPM. Dr. Jinxing Ji is a consulting metallurgist with JJ Metallurgical Services. He is a registered professional engineer (P.Eng.) in the province of British Columbia, Canada.

The PEA metallurgical testwork program between 2024 and 2025 was built upon the earlier metallurgical testwork in 2022 and 2023. The PEA metallurgical testwork program focused on sequential selective flotation for the silver zone to produce a silver/lead concentrate and a zinc concentrate, and on bulk flotation and cyanide leach for the gold zone. Five composite samples from the silver zone and one composite sample from the gold zone were tested. These composite samples were prepared from a large number of drill holes and intervals across the entire deposit.

The QP reviewed the selection of these composite samples, metallurgical testing procedures and data interpretations, and considered they are appropriate and reasonable. Nevertheless, the completed metallurgical testwork programs are still preliminary and limited to represent the entire deposit and the intended flowsheets. Further metallurgical testwork programs have been planned.

25.7.1 Metallurgical Composite Samples

A large number of drill holes and intervals were chosen to prepare the composite samples used in the PEA metallurgical testwork program. In the silver zone, three domain composite samples and two LOM composite samples were prepared and tested for flotation.

· The Oxidized domain sample was prepared with 112 intervals in 46 drill holes with the targets of 61 g/t<br>silver, 0.44% lead, 0.08% zinc and 0.20% sulfur.
· The Transitional domain sample was composited from 108 intervals in 45 drill holes with the grades expected<br>to be 61 g/t silver, 0.48% lead, 0.65% zinc and 0.89% sulfur.
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· The Fresh (Sulfide) domain sample contained 107 intervals in 55 drill holes to target 59 g/t silver, 0.42%<br>lead, 0.89% zinc and 1.75% sulfur.
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· The first LOM composite sample with silver grade targeted at 60 g/t consisted of 12.5% Oxidized domain,<br>2.5% Transitional domain and 85.0% Sulfide domain.
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· The second LOM composite sample with silver grade targeted at 74 g/t was made up of 17.0% Oxidized domain,<br>3.0% Transitional domain and 80.0% Sulfide domain.
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· In the gold zone, one composite sample was prepared for the testing of gravity concentration, flotation<br>and cyanide leach. This composite sample contained 89 intervals in 23 drill holes with gold grade targeted at 1.01 g/t.
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25.7.2 Comminution
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The comminution testing was carried out with three composite samples from the upper silver zone, lower silver zone and lower gold zone. Each sample contained 20–23 intervals from three drill holes. The measured rod mill work index, ball mill work index and abrasion index values were 10.1–12.3 kWh/t, 10.7–12.8 kWh/t and 0.038–0.075 g, respectively. These values indicate that these three samples were moderately hard and mildly abrasive.

25.7.3 Flotation of the Mineralized Materials in the Silver Zone

Sequential selective flotation was investigated and proven to be successful to produce two separate marketable concentrates for the mineralized materials in the silver zone. For the LOM composite sample which was blended at a ratio of 12.5% Oxidized domain, 2.5% Transitional domain and 85.0% Sulfide domain with head grades of 54 g/t silver, 0.39% lead and 0.66% zinc, the silver/lead concentrate contained 3,675 g/t silver and 24.4% lead with 82.1% silver recovery and 74.1% lead recovery, and the zinc concentrate contained 332 g/t silver and 45.3% zinc with 6.8% silver recovery and 69.3% zinc recovery.

25.7.4 Metallurgical Testing of the Mineralized Materials in the Gold Zone

A composite sample (1.01 g/t gold, 11 g/t silver, 0.060% copper and 3.07% sulfur) in the gold zone was subjected to gravity concentration, flotation and cyanide leach. About 45% gold recovery is expected from a commercial gravity concentration circuit. The whole-ore cyanide leach resulted in 94.0% gold recovery and 50.8% silver recovery. This composite sample was amenable to bulk flotation with 98.0% gold recovery and 94.7% silver recovery at 10.9% mass pull. The cyanide leach of the bulk flotation concentrate achieved 95.3% gold recovery and 63.8% silver recovery. A selective flotation test generated a copper/gold concentrate which contained 17.2% copper, 370 g/t gold and 1,564 g/t silver with 38.7% copper recovery, 51.5% gold recovery and 22.8% silver recovery.

25.7.5 Process Flowsheet for the Mineralized Materials in the Silver Zone

A conventional process flowsheet is applicable. The process plant starts with a primary crushing circuit. The crushed material is then conveyed to a stockpile. After being reclaimed, the crushed material is conveyed to a grinding circuit, which consists of a SAG mill, a ball mill and a pebble crusher. Lime and zinc sulfate are added to the grinding circuit. The screen undersize of the SAG mill product is combined in a pumpbox with the ball mill product, and then pumped to a cyclone for classification. The cyclone underflow returns to the ball mill for further size reduction, and the cyclone overflow flows by gravity to a trash screen. The screen undersize then flows to a conditioning tank in the silver/lead flotation circuit.

Collectors AP3418A and A404 are used to float the silver/lead mineralization. The silver/lead are first floated in a rougher circuit. After regrinding, the rougher concentrate is conditioned and then upgraded in a three-stage cleaner circuit. The first stage cleaner is operated in an open circuit to reject zinc and pyrite. The final silver/lead concentrate is thickened and then filtered.

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The tailings from the silver/lead circuit will be thickened first, the thickener underflow is then forwarded to a conditioning tank in the zinc flotation circuit. Dilution water will be added as required to a targeted pulp density. Lime will be added first to raise the pH. Then, copper sulfate will be added to activate the zinc. The activated zinc is then recovered by flotation using either a SIPX or AP3418A collector. After regrinding, the zinc rougher concentrate is conditioned again and then upgraded in a three-stage cleaner circuit. The final zinc concentrate is thickened and then filtered. The first stage cleaner circuit will be operated in an open circuit or in a closed circuit, subject to the zinc recovery and pyrite rejection.

The tailings from the zinc circuit will be thickened and the resultant thickener underflow will be pumped to the TSF for disposal.

25.7.6 Process Flowsheet for the Mineralized Materials in the Gold Zone

The crushing and primary grinding circuits are similar to the mineralized materials in the silver zone. However, a gravity concentration circuit will be included in the grinding circuit to treat a portion of cyclone underflow. The resultant gravity concentrate will be subjected to intensive cyanide leach to dissolve gold and silver. The pregnant solution from the intensive cyanide leach will be forwarded to an electrowinning circuit to reduce the dissolved gold and silver into metallic gold and silver.

The cyclone overflow flows by gravity to a trash screen. The screen undersize then flows by gravity to a conditioning tank in the gold flotation circuit. Collectors SIPX and AF208 are used to float the gold/silver mineralization. Only rougher and scavenger flotation are required. After regrinding, the flotation concentrate will be subjected to cyanide leach. The leached slurry is then forwarded to an CCD circuit. The overflow (pregnant solution) from the CCD circuit is filtered to remove all suspended solid. The filtrate is then deaerated and then the dissolved gold and silver are reduced to metallic gold and silver by zinc dust. The precipitate is filtered, collected, dried and smelted to produce the gold/silver dore.

25.8 Mineral Resource Estimate

The Mineral Resource Estimate for the Carangas Project is based on drilling and geological work completed up to 2023 and outlines a large epithermal system with silver-rich mineralization near surface and a deeper gold-dominant zone. The available drilling, sampling, and geological interpretation provide enough information to support the current Mineral Resource estimate and its use in the PEA. For the current update, no new drilling was added; instead, the resource was updated by revising NSR values using updated metal prices and technical assumptions. A new pit shell was generated using these inputs, and the Mineral Resources reported in this study replace all previously published resource figures.

The Mineral Resource consists of Indicated and Inferred categories only, as the current drilling density and geological confidence are not sufficient to define any Measured Mineral Resources. Some uncertainty remains regarding the exact shape and true thickness of the mineralized zones, because mineralization orientation is not fully constrained and reported intercepts represent down-hole lengths. Despite these limitations, the Mineral Resource estimate is considered suitable for a PEA-level evaluation. Further drilling, improved structural information, and refinement of the geological model will be required to increase confidence, support potential resource upgrades, and advance the project toward more detailed technical and economic studies.

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Under the assumptions presented in this Technical Report, and based on the data available as of February 1, 2026, the QP estimated the Mineral Resources of the Project meet the 2014 CIM Definition Standards, the 2019 CIM Best Practice Guidelines, NI 43-101 guidelines and show RPEEE.

25.9 Mining Methods

Reasonable open pit mine plans, mine production schedules, and mine capital and operating cost estimates have been developed for the Carangas project PEA, mining 250 Mt of silver zone resource containing 36 g/t silver, 0.34% lead, and 0.64% zinc, 51 Mt of gold zone resources containing 12 g/t silver and 0.74 g/t gold, all at a 1.8 waste to mill feed mining ratio.

Pit and stockpile layouts and mine operation plans are typical of other regional open pit metal mines. Contractor-managed open pit activities have been proven effective in these other regional operations.

The mine plan and estimated mine capital and operating cost estimates are reasonable at a scoping level of engineering and support the cash flow model and financials developed for the PEA.

25.10 Recovery Plan

The recovery methods align with conventional base-metal practices in the industry. Comminution, flotation recovery of payable metals, handling of tailings, leaching and gold recovery are achieved through typical processes that are commonly used in the industry for similar projects. Testwork results were used to develop the resulting flowsheet suitable for each stage over the life of mine.

The recovery methods utilize a staged expansion approach to appropriately manage varying silver, lead, zinc and gold grades throughout the life of mine without incurring excessive capital costs early in the project. The expansions utilize twinned or parallel equipment wherever possible, as well as de-risked brownfield expansion activities and simplified engineering.

25.11 Infrastructure

The infrastructure design is based on reasonable assumptions and is suitable for this level of study. The Company has engaged with local governments and businesses to establish the inputs used to support the Project.

The Project site is currently accessible by vehicle by driving approximately 200 km via RN 12 from the City of Oruro to the Town of Sabaya. From Sabaya, a 35 km gravel road leads to the Carangas site. Employees, fuel, reagents, supplies, and concentrate will be transported to project via RN 12 either from the City of Oruro or from Chile.

The concentrate products will be loaded on bulk haulage highway trucks and transported via the access road to Sabaya. The bulk haulers will then follow Bolivian RN 12, Chilean RN 15, Chilean RN 5, Chilean RN 16 to the port of Iquique. The bulk concentrate will then be unloaded, stockpiled, assayed, and loaded on ships.

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The on-site roads, process plant building, support buildings, and stockpiles have been developed at a conceptual level consistent with a PEA.

The TSF and WRSF have been designed in an area that has not been previously studied. As a result, no geotechnical data is currently available for the region. The design standards for the TSF and WRSF are based on the relevant national and international guidelines for the construction and operation of mining tailings storage facilities in Bolivia.

25.12 Markets and Contracts

No market studies or product valuations were completed as part of this study. Market price assumptions were based on a review of public information, industry consensus, standard practice, and specific information from comparable operations.

Silver-lead and zinc concentrates are widely traded and can be marketed directly from producer to smelter or via third-party concentrate trading entities. It is assumed that the concentrate contains negligible deleterious elements that would impact marketability.

The market for gold doré is widely traded and can be marketed domestically or internationally with significant optionality regarding the final customer. It is assumed that the doré contains negligible deleterious elements that would impact marketability.

A marketing study was not conducted to determine indicative treatment and refining terms. Marketing, refining, and transportation costs, along with payability terms, were informed by preliminary discussions with a major commodity trading company and information from comparable recent studies.

25.13 Environmental, Permitting and Social Considerations

Select socio-environmental baseline and supporting studies were completed by Tierralta on behalf of NPM. These studies were completed in 2024 and include work in the following areas: watershed characteristics; hydrology and meteorology; groundwater and surface water quality; wetlands; biodiversity of flora; and ambient air quality. Social and community studies consisted of a review of secondary data sources and focused on the following areas: population; migration and immigration; education; health; energy use, and economic activity.

Key findings from the preliminary baseline studies conducted in 2024 indicated the following key findings:

● The<br> Carangas micro-basin was characterized as elongated, low-slope terrain with limited channel<br> networks and low water regulation capabilities.
● The<br> region faces historical water shortages from September to November.
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● Groundwater<br> and surface water are generally of good quality for most parameters but report consistently<br> elevated chemical/biological oxygen demands and high ammonia levels.
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● Ecological<br> surveys identified locally high diversity of flora species on slopes and within localized<br> lotic and wetland ecosystems, which are critical for ecosystem health and local land use<br> activities such as livestock raising.
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● The<br> area of influence for the Project holds approximately 1,130 residents. Socio-economic assessments<br> reveal vulnerability of the local population with high rates of poverty and indigence.
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The scope of the 2024 baseline studies is appropriate for the current design and permitting phase of the Project and forms a strong base for planning and execution of future studies. As the Project moves into the feasibility design phase and as the environmental assessment/permitting process progresses, there are identified gaps and additional field and desktop studies will be required.

A comprehensive system of socio-environmental monitoring and management plans will be developed based on applicable regulatory requirements and considering community concerns and international best practice. These plans will ensure that the Project operates in a manner that protects Project personnel, the environment, and local communities, while promoting positive impacts. Environmental monitoring for mining projects in Bolivia is regulated by the MMAyA, through national and departmental environmental authorities, under the General Environmental Law and by local regulations. Post-closure environmental monitoring will align with the requirements for mine closure in Bolivia.

During 2025, NPM initiated and advanced the regulatory transition process to convert the PELs to Administrative Mining Contracts (AMCs). Work and Investment Plans for the Granville and Colapso mining areas were prepared and submitted to the relevant authorities and subsequently reviewed with SERGEOMIN and AJAM as part of the AMC evaluation and approval process.

Upon approval of the AMCs by AJAM, the Company intends to apply for environmental categorization and initiate preparation of the Environmental Impact Assessment (EEIA) required by future project environmental permitting for mine development. As part of the permitting stage, the Project will submit a Closure Plan along with the EEIA application. The Plan will include objectives, closure program details, rehabilitation measures, contaminant control measures, and post-closure monitoring. The Closure Plan will be prepared in parallel with mine facility designs, incorporate baseline information studies, and other operational and long-term planning efforts.

Throughout 2024, 2025, and early 2026, NPM implemented an extensive community engagement program designed to strengthen relationships with local stakeholders and facilitate the advancement of the Project. Engagement activities included formal and informal meetings with community members and authorities, technical workshops, educational outreach programs, stakeholder mapping, sociocultural baseline studies, and communication campaigns tailored to local customs and traditions. Additional support initiatives included veterinary assistance and technical support programs for local camelid livestock producers, as well as participation in civic, social, and cultural activities within the community.

On January 26, 2026, NPM and the Carangas Community executed a formal Framework Agreement regulating the relationship between the Project and the Community. This agreement represents the most significant advancement achieved to date in obtaining the social permits required for project development and establishes the principal governance instrument currently in force between both parties.

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25.14 Capital Cost Estimate
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The capital cost estimate was developed in Q2 2026 to target a level of accuracy of -30% to +50%, which aligns with an Association for the Advancement of Cost Engineering International (AACE International) Class 5 level estimate. The estimate includes mining, processing, on-site infrastructure, off-site infrastructure, project indirects, project delivery, owners’ costs, and provisions. The total initial capital costs for the Project are estimated at US$644.5 million, including US$55.3 million of capitalized operating costs, and US$110.3 million of contingency. The total expansion capital costs are US$422.7 million and the total LOM sustaining costs are estimated at US$166.5 million, while the closure costs are estimated at US$149.8 million.

25.15 Operating Cost Estimate

The total operating costs for the Project are estimated at US$16.93/t or US$4,258.4 million over the 19-year mine life. These operating costs do not include pre-production operating costs. This estimate considers the initial development of a process plant designed to treat 22,000 t/d (8.0 Mt/a) of mineralized material with the expansion phase designed to treat 44,000 t/d (16.0 Mt/a). Process unit operations were benchmarked against similar or comparable processing plants to ensure accuracy of cost estimates.

25.16 Economic Analysis

An economic model was developed to estimate the project's annual pre-tax and post-tax cash flows, sensitivities, and net present value results using a 5% discount rate. Based on the assumptions and parameters, the PEA shows positive post-tax economics of US$2.65 billion NPV5% and 35.9% post-tax IRR.

A sensitivity analysis was conducted on the base-case pre-tax and post-tax NPV and IRR of the project using the following variables: metal prices, discount rate, operating costs, initial capital expenditure, metal recovery, and head grade. The Project is most sensitive to changes in recovery, commodity price and head grade, and less sensitive to changes in operating costs and initial capital costs.

The PEA is preliminary in nature, that it includes inferred mineral resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as mineral reserves, and there is no certainty that the preliminary economic assessment will be realized.

25.17 Risks and Opportunities
25.17.1 Introduction
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The risks and uncertainties identified below could influence the economic viability of the Carangas Project if not addressed. Reduced confidence in grade continuity or mineralization geometry may limit the extent to which Mineral Resources can be upgraded or converted to Mineral Reserves in future studies. Unforeseen geological complexity or lower-than-expected continuity may require additional drilling and associated costs. Non-technical risks, such as changes in surface access, permitting requirements, or fiscal terms, may introduce delays or alter economic outcomes.

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Conversely, successful drilling of open extensions or IP anomalies could increase the overall Mineral Resource base and enhance the Project’s economic potential. Improvements in structural understanding, metallurgical performance, or pit-slope design could also strengthen future project economics. Overall, while risks exist and are typical for a project at the PEA stage, there remain clear opportunities to improve confidence and support the Project’s continued development.

25.17.2 Risks
25.17.2.1 Exploration
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Several factors could affect the reliability of the exploration information and the confidence in the current Mineral Resource estimate for the Carangas Project. The most significant technical risks are mentioned below:

● Uncertain<br> orientation and geometry of mineralization, due to limited structural data and lack of oriented<br> drilling. This creates uncertainty in true thickness and continuity.
● Down-hole<br> intervals may not represent true thickness, increasing uncertainty in grade distribution<br> and modelling confidence
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● Epithermal<br> system heterogeneity, including abrupt changes in grade, alteration, breccia development,<br> and metal zoning, may impact the predictability of the deposit.
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● No<br> new drilling or geological updates since the 2023 inaugural Mineral Resource, meaning the<br> current model relies on older datasets that may not fully capture system complexity.
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● Limited<br> drilling at depth, especially in the gold-dominant<br> zone, increases uncertainty regarding continuity beyond the conceptual pit.
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● Geophysical<br> chargeability anomalies remain untested, meaning their mineral potential is unknown and could<br> represent exploration risk if they do not contain economic mineralization.
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25.17.2.2 Drilling
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Key drilling risks relate to the following:

● Uncertain<br> orientation and geometry of mineralization, due to limited structural data and lack of oriented<br> drilling. This creates uncertainty in true thickness and continuity.
● Geological<br> complexity of the epithermal system where mineralization occurs across multiple lithologies,<br> including volcanic units, breccias, and dome-related<br> intrusive rocks, which introduces variability in geometry and grade continuity that may not<br> be fully captured by the current drill spacing.
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● All<br> drilling data supporting the Mineral Resource were completed between 2021 and early 2023,<br> with no subsequent drilling. Any changes in geological understanding will require new drilling<br> to validate current interpretations.
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25.17.2.3 Sample Preparation
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Key risks related to sample preparation and analysis are as follows:

● Analytical<br> variability at very high grades where silver and gold occur locally at very high grades,<br> which increases the risk of analytical variability, nugget effects, or over-limit<br> assay issues, particularly for gold where free gold may be present.
● Potential<br> for contamination during sample handling. Despite strong blank performance, there is always<br> a residual risk of contamination during crushing, pulverizing, or analytical stages, especially<br> in high-throughput laboratory<br> environments.
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25.17.2.4 Metallurgical Testing
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Key metallurgical testing risks relate to the following:

● The<br> rejection of zinc in the silver/lead flotation circuit is adversely impacted by the oxidized<br> materials.
● The<br> Deleterious elements in the silver/lead concentrate have not been assayed,
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● The<br> cyanide in the silver/lead flotation circuit reduces silver recovery.
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● The<br> slurry viscosity in the zinc rougher flotation circuit is excessively high with lime addition<br> to reach pH over 11.0. To mitigate this high slurry viscosity issue, the lime addition in<br> the rougher circuit is capped at 500 g/t Ca(OH)2.
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● The<br> high-level copper in the gold flotation concentrate may negatively impact cyanide leach performance<br> when the gold flotation concentrate is subjected to cyanide leach.
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● The<br> dissolved copper in the pregnant leach solution may precipitate together with gold and silver<br> in the Merrill Crowe circuit.
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● The<br> mineralized samples from the silver zone, which were used in the flotation testwork, targeted<br> silver grade at 60 g/t or higher. The 2026 mine production schedule shows a LOM average silver<br> grade of 36 g/t. The mineralized materials with 36 g/t silver or lower have not been tested<br> yet.
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The mineralized samples from the gold zone, which were used in the metallurgical testwork, targeted gold grade at 1.00 g/t or higher. The 2026 mine production schedule shows a life-of-mine average gold head grade of 0.74 g/t. The mineralized materials with 0.74 g/t Au or lower have not been tested yet.

25.17.2.5 Mineral Resource Estimate

Key risks associated to the Mineral Resource Estimate are as follows:

● Reliance<br> on drilling completed up to 2023 to estimate the current Mineral Resource, with no new drilling<br> or geological reinterpretation incorporated since that time.
● Limited<br> structural control on mineralization geometry.
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● Grade<br> variability inherent to epithermal systems.
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● Dependence<br> on modeling assumptions related to grade capping, compositing length, domain boundaries,<br> interpolation method (ID2), and search parameters, all of which may change with additional<br> data.
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● Sensitivity<br> to economic parameters for the constrained conceptual open-pit<br> shell and cut-off grade based<br> on metal prices, recoveries, and cost assumptions; changes in these inputs could materially<br> affect reported tonnage and grade.
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● Absence<br> of Measured Mineral Resources as the current drill spacing and geological confidence did<br> not support the Measured classes.
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● Mineralization<br> below the conceptual pit and within the Lower Copper Zone remains insufficiently drilled<br> and is not included in the Mineral Resource estimate.
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● Approximately<br> 1.85% of the Mineral Resources included in the PEA pit design are located within adjacent<br> Bolivian State-held concessions that are not currently controlled by the Company; failure<br> to secure access or a mining agreement could require pit redesign and may negatively impact<br> the PEA economic outcomes.
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25.17.2.6 Mining
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For risks associated with mining methods, refer to Section 16.10.

25.17.2.7 Recovery Methods

The study was completed based on limited metallurgical testwork. As a result, certain process design assumptions remain to be confirmed through future testwork and engineering studies. The key areas of uncertainty and potential design impact are summarized below:

● The<br> grinding circuit design is based on assumed drop-weight test Axb values provided by the client,<br> as no coarse material competency testwork has been completed to date. Although the SAG mill<br> design includes additional design margin to account for this uncertainty, there is a risk<br> that the grinding circuit may be undersized if future comminution testwork indicates that<br> the mineralized material is more competent than currently assumed.
● In<br> general, grinding equipment has been selected based on the comminution test data available<br> at the time of the study. If future testwork identifies mineralized material hardness values<br> greater than the design values used, selected equipment may be undersized, potentially affecting<br> throughput.
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● The<br> silver-lead-zinc flotation flowsheet has been developed based on the metallurgical testwork<br> completed to date. Given the limited dataset, the selected flowsheet may not represent the<br> optimum configuration for metallurgical performance, concentrate quality, reagent consumption,<br> or operating flexibility. Additional flotation testwork and flowsheet optimization studies<br> are recommended.
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● Gold<br> leaching testwork remains limited, and the current leach circuit design may not be fully<br> optimized. Additional testwork is required to confirm leach kinetics, achievable gold recovery,<br> reagent requirements, and the appropriate retention time for design.
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● Copper<br> dissolution observed during gold leaching testwork indicates the potential for elevated cyanide<br> consumption. Further testwork is recommended to quantify soluble copper behaviour, cyanide<br> demand, and potential mitigation strategies, as these factors may impact operating costs<br> and plant design requirements.
--- ---
● Concentrate<br> filtration moisture values have been assumed from benchmark data appropriate for this level<br> of study. Filtration testwork is required in future study phases to confirm achievable final<br> moisture contents, validate equipment sizing, and support concentrate handling, transport,<br> and operating cost assumptions.
--- ---
25.17.2.8 Infrastructure
--- ---

25.17.2.8.1            Hazard Considerations

Assumptions for seismic risk and geohazards for the project have been assumed for the PEA. However, in the future site-specific seismic hazard, geohazards, and climate-driven changes studies will need to be performed and if they are different from those assumed there could be an impact on the design of infrastructure, i.e. capital, sustaining capital, and operating cost.

Risks related to on-site infrastructure include:

● Lack<br> of geotechnical information throughout the project site could impact the stability and locations<br> of the proposed designs of infrastructure.
● Building<br> costs at the PEA-level are based on benchmark costs and do not reflect actual structurally<br> engineered designs. These costs could increase with a more detailed design that considers<br> specific site conditions design criteria.
--- ---
● There<br> is currently no geochemical data on the waste rock and tailings for the PEA. Geochemical<br> analysis are proposed for the next phase of the project. Depending on the results of this<br> program, the design and cost of the TSF and WRSF may increase.
--- ---

25.17.2.8.2            TSF and WRSF

● The<br> project has proposed a TSF and WRSF in an area that has not been previously studied. As a<br> result, no geotechnical data is currently available for the region. The key risks associated<br> with the TSF and WRSF locations and study area are as follows:
● Ground<br> conditions, geological containment, and slope stability within the proposed TSF and WRSF<br> footprints are unknown, as no geotechnical investigation program has been completed to date.
--- ---
● There<br> is potential for increased project costs if geotechnical or hydrogeological conditions differ<br> from the assumptions used in this study. Any deviations may affect capital costs, sustaining<br> capital, and operating costs associated with the TSF and WRSF.
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25.17.2.9 Environmental, Permitting and Social Considerations
--- ---

The Project is subject to several environmental, permitting, social, and community-related risks that could affect its development timeline, cost structure, or long-term viability. The key risks identified are as follows:

● Critical<br> Path Permitting – The Environmental Licence and associated water use, discharge, and<br> other associated authorizations are prerequisites for construction. Any delays in obtaining<br> or modifying these authorizations could defer project start-up.
● Regulatory<br> Uncertainties - Mining in Bolivia involves navigating a complex, highly regulated, and politically<br> dynamic legal landscape. While the government has made recent pushbacks to streamline processes<br> and court international investors, several core regulatory and institutional risks remain<br> critical for project development. Ongoing monitoring of the legal, regulatory and political<br> aspects are critical to the early identification and mitigation of potential future risks<br> to project approvals and controls.
--- ---
● Environmental<br> Sensitivities – The Project’s area of influence includes streams, localized riparian<br> areas and wetlands, and land used for seasonal agricultural purposes by the local community.<br> The local water basin shows low water regulation capacity with seasonal water shortages from<br> September to November. The utilization of local surface and groundwater resources to<br> supply the project will require careful analysis and management. Strict management of project<br> impacts and monitoring of these environmental sensitivities will be required.
--- ---
● Community<br> Acceptance: Local communities, including agricultural land users and small-scale miners,<br> may raise concerns related to land use, employment opportunities, or perceived environmental<br> impacts. In particular, there may be local concerns due to the deposit's proximity to local<br> infrastructure and the requirement to physically relocate the existing Carangas village.<br> Failure to address these concerns through ongoing engagement and benefit-sharing could affect<br> the social license to operate.
--- ---
● Post-Closure<br> Obligations – The need for long-term water quality monitoring and potential acid rock<br> drainage/metal leaching management may extend post-closure obligations and associated costs.
--- ---
25.17.3 Opportunities
--- ---
25.17.3.1 Exploration
--- ---
● Expand<br> the Mineral Resource through step-out<br> drilling, as gold mineralization remains open to the north, northeast, and at depth.
--- ---
● Upgrade<br> confidence in the current Mineral Resource by conducting infill drilling to improve continuity<br> and support conversion of Inferred to Indicated categories.
--- ---
● Test<br> multiple strong IP chargeability anomalies beyond the drilled area that show similar geophysical<br> signatures to the known mineralization and may host additional mineralized zones.
--- ---
● Improve<br> understanding of mineralization controls by acquiring oriented core and more detailed structural<br> data during future drilling.
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● Enhance<br> geological modelling by refining interpretations of breccia bodies, alteration patterns,<br> and intrusive contacts as more drill data are collected.
--- ---
● Increase<br> project value if exploration drilling confirms additional mineralized zones indicated by<br> geophysics or extends the high-value<br> deep gold system.
--- ---
25.17.3.2 Drilling
--- ---
● Complete<br> an infill drilling to improve confidence
--- ---
● Step-out<br> drilling at depth and along strike to confirm mineralization continuity
--- ---
● Use<br> of oriented core in future programs collecting oriented core and conducting detailed structural
--- ---
● Improve<br> geological and domain modeling through additional drilling to refine lithological, alteration,<br> and mineralization domains
--- ---
● Future<br> drilling programs can be seamlessly incorporated into the existing validated database structure
--- ---
● Improved<br> geological modeling with additional structural data
--- ---
● Potential<br> to upgrade resource confidence with additional infill drilling.
--- ---
25.17.3.3 Sample Preparation
--- ---
● Continue<br> the QA/QC protocols that follows industry benchmarks
--- ---
● Use<br> of multiple control sample types in the QA/QC protocols
--- ---
● Keep<br> the established sample security and chain of custody
--- ---
25.17.3.4 Metallurgical Testing
--- ---
● The<br> coarse assay sample rejects were used to prepare metallurgical samples for flotation testwork.<br> The flotation performance is sensitive to oxidation. If the fresh drill core intervals are<br> used, the flotation performance is expected to improve.
--- ---
● In<br> the zinc flotation circuit, the use of collector AP3418A has proven successful to replace<br> the xanthate SIPX. The collector AP3418A is more selective against pyrite.
--- ---
● The<br> silver zone contains 0.057 g/t gold on average based on the 2026 mine production schedule.<br> A portion of this gold may be recoverable by gravity concentration.
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● For<br> the mineralized materials in the gold zone, the selective flotation testing has proven to<br> produce a marketable high-grade copper/gold concentrate. In addition, a high-grade gold concentrate<br> with over 70 g/t gold may be produced as well. These two concentrates will increase the revenue<br> of copper, silver and perhaps gold as well. After majority of copper is removed, the bulk<br> flotation concentrate will be more amenable to cyanide leach in terms of cyanide consumption<br> and gold dore quality. If majority of copper and silver are removed by selective flotation<br> before bulk flotation, the carbon-in-pulp cyanide leach may be feasible for the bulk gold<br> flotation concentrate to replace the CCD and Merrill Crowe.
--- ---
● The<br> flotation performance may be improved when the regrinding is applied to the 1^st^<br> cleaner concentrate as opposed to the rougher concentrate.
--- ---
25.17.3.5 Mineral Resource Estimate
--- ---
● Future<br> drilling programs can be seamlessly incorporated into the existing validated database structure.
--- ---
● Improved<br> geological modeling with additional structural data.
--- ---
● Potential<br> to upgrade resource confidence with additional infill drilling.
--- ---
25.17.3.6 Recovery Methods
--- ---

There may be opportunities to optimize the process flowsheet for the Project following completion of suitable metallurgical testwork. Future studies should include engineering trade-off evaluations to confirm and refine key design assumptions, including the following:

● The<br> grinding circuit design is based on assumed drop-weight test Axb values provided by the client,<br> as no coarse material competency testwork has been completed to date. The SAG mill design<br> currently includes additional design margin to account for this uncertainty. Should future<br> comminution testwork indicate that the mineralized material is less competent than assumed,<br> there may be an opportunity to reduce the SAG mill size or installed power requirements.
● Additional<br> metallurgical testwork should be completed to confirm reagent consumption rates for the concentrator<br> flotation circuit. The results should be used to validate operating cost assumptions and<br> identify opportunities to optimize reagent addition strategies.
--- ---
● Further<br> leach testwork should be undertaken to confirm the required retention time for the gold plant.<br> If testwork demonstrates that target gold recoveries can be achieved at shorter retention<br> times, there may be an opportunity to reduce leach circuit volume and associated capital<br> and operating costs.
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25.17.3.7 Environmental, Permitting and Social Considerations
--- ---

In addition to the identified risks, the Carangas Project presents several environmental, permitting, social, and community-related opportunities that could positively influence Project execution and long-term performance:

● Expediting<br> Required Baseline Programs - Building on current information, prioritize and expedite well<br> informed environmental, socio-economic, and cultural baseline studies to support the Project<br> through feasibility design, permitting, development, operations and closure phases.
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● Collaboration<br> between Technical Teams - There are opportunities to work closely with ongoing geological<br> and geotechnical drilling program teams to efficiently collect required geochemical samples<br> and establish a hydrological and hydrogeological monitoring and testing program that will<br> be required for future feasibility and permitting phases of the Project.
--- ---
● Community<br> Development Partnerships – Opportunities exist to formalize benefit-sharing initiatives,<br> including local hiring, training programs, and procurement from local suppliers, strengthening<br> social license to operate.
--- ---
● Environmental Stewardship Leadership – Implementation<br>of best practice biodiversity management, progressive reclamation, and water stewardship programs can enhance the project’s reputation<br>and reduce regulatory and reputational risk.
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26 Recommendations
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26.1 Introduction
--- ---

The Carangas Project shows positive economics based on a conceptual level of design, as demonstrated by the results presented in this technical report. It is recommended that the project be advanced to a feasibility-level study. The following sections detail the recommended future work for the project. The estimated costs are summarized in Table 26-1.

Table 26-1:         RecommendedWork Program

Program Component Estimated Total Cost (US$M)
Drilling 12.0
Geology<br> and Exploration 0.7
Mining<br> and Mine Open Pit Geotechnical Program 3.8
Metallurgical<br> Testing 2.5
Recovery<br> Methods 3.5
Infrastructure<br> Geotechnical 1.4
TSF<br> and WRSF Analyses and Studies 0.6
Hydrogeological<br> - Stage 1 1.5
Hydrogeological<br> - Stage 2 2.0
Hydrological 1.0
Geochemical 2.5
Environmental 6.0
Total 37.5
26.2 Exploration and Mineral Resource
--- ---
26.2.1 Exploration and Drilling
--- ---

Based on the outcomes of the Mineral Resource estimate and the current stage of the Project, it is recommended that additional drilling be undertaken.

● Infill drilling: The existing drilling grid is largely 50 m by 50 m in the majority of the drilled<br> area and supports the Indicated Mineral Resources category. An appropriate amount of infill<br> drilling is needed in the core area of the known mineralization system to further confirm<br> the continuity of mineralization, thereby enhancing confidence in the Mineral Resources and<br> facilitating future advanced technical and economic studies of the Project.
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● Step-out drilling: Gold mineralization continues below the conceptual pit constraint and remains<br> open to the north and northeast directions. Therefore, step-out drilling is justified to<br> unveil the potential of additional Mineral Resources.
--- ---
● Exploration drilling: Multiple strong IP chargeability anomalies were identified beyond the drilled<br> areas. These anomalies displayed geophysical signatures similar to those of the known mineralization<br> system. It is reasonable to anticipate that these anomalies may host mineralization similar<br> to that drilled so far and to hope for the addition of Mineral Resources through future drilling<br> campaigns.
--- ---
● Data<br> verification:
--- ---
o Maintain<br> the current data verification protocols for all future drilling and exploration programs.
--- ---
o Conduct<br> data verification concurrently with future drilling campaigns to ensure timely identification<br> and resolution of any data issues.
--- ---
o Include<br> oriented core drilling and enhanced structural logging in future programs to improve confidence<br> in mineralization geometry and true thickness interpretation.
--- ---
o Update<br> and/or re-validate the topographic<br> surface and survey data as new drilling or infrastructure is added to the project.
--- ---
o Ensure<br> that all new data is integrated into the existing database.
--- ---
26.2.2 Geology Study, Mapping, and Prospecting
--- ---
● Maintain<br> the partnership with the universities in Bolivia to continue geological studies on the Carangas<br> deposit to further understand the mineralization styles and genesis, and support future exploration<br> targeting.
--- ---
● Initiate<br> exploration programs including geological mapping and prospecting over the IP chargeability<br> anomalies for refining targets of drilling tests.
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26.2.3 Estimated Budget for Recommendations
--- ---

The estimated budget to complete the main activities recommended above is US$12.7 million, as outlined in Table 26-2.

Table 26-2         EstimatedBudget for Exploration and Mineral Resource Recommendations

Area Estimated Budget (US$M)
Geology and Mineral Resources
Drilling<br> 30,000 m 12.0
Geology<br> Study 0.3
Geological<br> Mapping and Prospecting 0.4
Total Estimated Cost – Next Phase of Study 12.7
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26.3 Mineral Processing and Metallurgical Testing
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The PEA metallurgical testing program has demonstrated that it was successful to produce marketable silver/lead concentrate and zinc concentrate by sequential selective flotation of the mineralized materials for the silver zone and to produce gold doré by cyanide leach of the bulk gold flotation concentrate for the mineralized materials in the gold zone. Further metallurgical testing is recommended to improve flotation and cyanide leach performance, finalize the flowsheets and generate necessary parameters for the process plant design:

● Coarse<br> assay sample rejects were used in the flotation testwork. These materials might have been<br> moderately oxidized during storage. The flotation performance is sensitive to oxidation.<br> Therefore, a composite sample, which consists of fresh core intervals in the silver zone,<br> is recommended for further flotation optimization.
● The<br> 2026 mine production schedule shows lower head grades compared with the 2024 mine production<br> schedule. There is a need to test the low-grade materials in the future.
--- ---
● The<br> 2026 mine production schedule shows 0.057 g/t Au on average in the silver zone. The gravity<br> concentration may be able to recover a portion of this gold. Therefore, the gravity concentration<br> testing is recommended.
--- ---
● Besides<br> silver, lead, zinc, iron and sulfur in the silver/lead concentrate and zinc concentrate,<br> the contents of gold, arsenic, antimony, bismuth, mercury, tin, chloride, fluoride, and silica, etc.,<br> need to be analyzed for these two concentrates.
--- ---
● Three<br> composite samples were subjected to the measurements of rod mill work index, ball mill work<br> index and abrasion index. More samples of the mineralized materials from different locations<br> in the deposit are recommended for comminution testing. In addition to the rod mill work<br> index, ball mill work index and abrasion index, the crusher work index, drop-weight testing<br> and SMC testing are recommended.
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● Some<br> of zinc minerals may have been activated in-situ in the deposit. Mineralogical investigations<br> and detailed chemical analyses are recommended to identify whether any elements trigger this<br> natural activation. Furthermore, it is recommended to investigate the surface conditioning<br> to de-activate zinc minerals in the silver/lead flotation circuit.
--- ---
● High<br> levels of clay minerals are present in the silver zone. Mineralogical examinations will be<br> carried out to determine the nature of clay minerals (such as, kaolinite, illite, halloysite<br> and chlorite, etc.) with close attentions to the swelling clay minerals.
--- ---
● Blending<br> and slurry viscosity modifier will be investigated to mitigate the negative impact of clay<br> minerals on flotation performance. Continue investigation of zinc rougher flotation at a<br> relatively low pH when the content of clay minerals is high.
--- ---
● The<br> operating conditions and chemicals in the silver/lead circuit are different from those in<br> the zinc circuit. The cross contamination of process water in these two flotation circuits<br> is expected in the future commercial operations. It is recommended to investigate the impact<br> of such cross contamination on the flotation performance in the silver/lead circuit and in<br> the zinc circuit.
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● For<br> the mineralized materials in the silver zone, thickening tests and slurry rheology measurements<br> are recommended for the flotation tailing, silver/lead concentrate and zinc concentrate.
--- ---
● For<br> the mineralized materials in the silver zone, filtration and transportable moisture limit<br> are recommended for the silver/lead concentrate and zinc concentrate.
--- ---
● For<br> the mineralized materials in the gold zone, further selective flotation testwork is recommended<br> to produce a marketable copper concentrate and a marketable gold concentrate to improve the<br> project economics and in the same time, to mitigate the negative issue of copper in the cyanide<br> leach of the bulk gold flotation concentrate.
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● The<br> regrinding of the bulk gold flotation concentrate improved gold recovery and silver recovery<br> in cyanide leach. However, more copper is expected to dissolve after regrinding. A series<br> of regrinding and cyanide leach tests are recommended to determine the optimum regrind size.
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● Dissolved<br> copper is present in the pregnant leach solution after cyanide leach of the bulk gold flotation<br> concentrate. Testwork is recommended to investigate the Merrill Crowe process with the objective<br> of minimizing copper co-precipitation while dissolved gold and silver are completely reduced<br> by zinc dust.
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● The<br> buildup of dissolved copper in the process water is less if CIP is applied instead of Merrill<br> Crowe. To demonstrate if the CIP is effective, the loading isotherm of gold, silver and copper<br> on the activated carbon is recommended using a pregnant leach solution from cyanide leach<br> of the bulk gold flotation concentrate. The first bulk flotation concentrate will be generated<br> without prior removal of a portion of copper by selective flotation, and the second bulk<br> flotation concentrate will be generated after a portion of copper is removed by selective<br> flotation.
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● For<br> the mineralized materials in the gold zone, thickening tests and slurry rheology measurements<br> are recommended for the flotation tailing and the cyanide-leached bulk flotation concentrate.
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● Acid<br> base accounting (ABA) and net acid generation (NAG) are recommended to determine the potential<br> of acid generation from the flotation tailings, cyanide leach residues and mine waste rocks.<br> For a few selected samples, additional tests are recommended, including detailed chemical<br> analysis, TCLP, SPLP, column leach and humidity cell testing.
--- ---
● The<br> raw water from the mine site will be collected for use in the flotation testwork of silver/lead<br> concentrate, zinc concentrate, copper/gold concentrate and gold concentrate. Flotation performance<br> will be compared between the raw water from the mine site and the municipal tap water.
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● A<br> mixed tailing water will be generated by combining the process waters from the silver/lead/zinc<br> flotation tailing, the gold flotation tailing and the detoxed cyanide leached gold flotation<br> concentrate. This mixed tailing water is then used in flotation of the silver/lead concentrate,<br> zinc concentrate, copper/gold concentrate and gold concentrate to investigate the impact<br> of residual cyanide, copper and zinc on the flotation performance.
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● The<br> estimated cost for this testwork is approximately US$2.5 million.
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26.4 Mineral Resource Estimate
--- ---

Further work is recommended to improve confidence in the Mineral Resource estimate and to support future upgrades in classification. The current resource model relies on drilling completed up to 2023, and no new geological or drilling data have been incorporated since the inaugural Mineral Resource. Additional drilling, particularly oriented core and closer-spaced infill holes, is recommended to better define the geometry of the mineralized zones and improve understanding of structural controls. These data will help reduce uncertainty in true thickness, refine mineralization domains, and support the conversion of Inferred Resources to higher-confidence categories.

The QP also recommends that future drilling programs expand coverage into areas where the deeper gold-dominant system remains open. Targeted step-out drilling may identify additional mineralized extensions, potentially increasing the resource footprint. Continued refinement of the geological model, incorporating updated interpretations of lithology, alteration, and breccias, will strengthen the basis for future resource evaluations.

Additional work on supporting datasets is also advised. This includes further metallurgical testwork to validate and refine recovery assumptions, geotechnical and hydrogeological investigations to improve slope and pit design constraints, and updated NSR modelling as new cost and process parameters become available. Continued engagement with local communities is essential to secure access for drilling and to ensure uninterrupted progress on future resource definition programs.

26.5 Mining
26.5.1 Mine Engineering
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The following recommendations are made with regards to advancing the mine engineering of the Carangas project to a Feasibility Study, with a budget for each recommended program included:

● Infill<br> exploration drilling with the aim of upgrading Inferred class Mineral Resources to Indicated<br> class Mineral Resources. (costed elsewhere in recommendations).
● Further<br> hydrogeological and hydrological characterization are required in the pit areas (costed elsewhere<br> in recommendations).
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● Geochemical<br> characterization of waste rock for the purposes of updated PAG modelling. It is possible<br> to utilize existing and planned exploration and geotechnical drill core for geochemical samples,<br> and no additional drilling has been planned for these studies in the estimated budget. (US$0.4<br> million).
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● Topsoil<br> and overburden assessment for the open pit to estimate topsoil and overburden storage requirements<br> (US$0.2 million).
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● Condemnation<br> drilling of the footprints identified for the waste rock storage facilities, as well as site<br> infrastructure. (US$1.0 million).
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● Drill<br> penetration and blast fragmentation studies, testing properties in all lithologies, as well<br> as within mineralized areas and within waste rock. It is possible to utilize existing and<br> planned exploration and geotechnical drill core for rock samples, and no additional drilling<br> has been planned for these studies in the estimated budget. (US$0.05 million).
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● Updates<br> to designs of open pits, waste storage piles, stockpiles, and mine haul roads incorporating<br> results from all other recommended work programs (US$0.15 million).
--- ---
● Including<br> further engagement with multiple potential contractors for mining operations, with more detailed<br> quotations for operations, emphasizing detailed operational quotations (US$0.05 million).
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A budget of US$1.9 million is estimated for the above work programs and studies.

26.5.2 Open Pit Geotechnical Program
26.5.2.1 Mine Geotechnical Drilling
--- ---

As part of the PEA, Ausenco looked at the development of an open pit geotechnical program. Targeted open pit geotechnical drilling using triple-tube HQ holes and televiewer with oriented cores. Recommend five drillholes on all sides of the planned open pit and an estimated total length of 3,500 m. Installation of vibrating-wire piezometers in selected holes. (US$1.5 million). All holes are to be televiewer logged and core should be geotechnically logged at the drill sites. Strength profiling by point load testing should also be continued, as per the current practice at the Project. The televiewer logging and standby costs are estimated to cost approximately US$80,000.

26.5.2.2 Surficial Mapping

Surficial mapping should be performed in the next phase to augment geotechnical information for the development of the open pit slopes. The mapping estimate is approximately US$10,000.

26.5.2.3 Laboratory Testing

Laboratory testing, including moisture content, gradation, triaxial strength, Atterberg limits, and density, should be performed to classify the alluvium. Also, the site-specific relationship between point load strength and unconfined compressive strength should be refined, and shear strength tests should be conducted to determine the basic friction angle for major rock types. The estimated cost for the laboratory testing is approximately US$15,000.

26.5.2.4 Structural Modelling

The structural model should be updated to incorporate all new drilling available at the time the PFS is completed. Borehole fault intersections should be traced for continuity to determine if structural control is present in the pit walls of either pit. The cost for updating the model is approximately US$60,000. The cost for reviewing the model is estimated to be approximately US$10,000.

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26.5.2.5 Rock Engineering Analyses
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After the new data is available, the rock engineering analysis should include:

● Update<br> the structural fabric database to incorporate both the open pit areas Evaluate the data for<br> spatial variation and/or lithologic correlations.
● Update<br> the rock mass geomechanical parameters and revise design values, as required.
--- ---
● Update<br> the alluvial geomechanical parameters and the spatial distribution of the alluvium thickness.<br> Evaluate hazard posed by upslope alluvial deposits relative to pit crests and design conceptual<br> mitigation measures accordingly.
--- ---
● Perform<br> slope stability analyses to optimize pit-wall design.
--- ---

The estimated cost for the rock engineering analysis is approximately US$90,000.

The estimated cost for the rock engineering analysis, including the expert review board, is approximately US$150,000.

A budget of US$1.9 million is estimated for the mine open pit geotechnical program.

26.6 Recovery Methods

The following activities are recommended to support the advancement of the project to the next phase:

● Explore<br> potential communition optimisation based on coarse material competency test work.
● Use<br> variability samples to identify opportunities for optimised equipment selection and sizing<br> related back to the mineral resource and mine plan.
--- ---
● Trade<br> off studies targeting NPV and IRR improvement scenarios.
--- ---
● Process<br> plant engineering, through criterion, lists, drawings, MTOs and cost estimates.
--- ---
● Cost<br> estimating.
--- ---
● Project<br> planning.
--- ---
● Technical<br> report support.
--- ---

The estimated costs for this work are US$3.5 million.

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26.7 Infrastructure
--- ---
26.7.1 Infrastructure Geotechnical
--- ---

Geotechnical investigations are required within the footprint of each planned structure, including the open pit. Table 26-3 specifies the number of field investigations required for each footprint. These investigations should be strategically located inside the footprint, applying engineering judgment, and considering the site’s topography and the design characteristics.

Table 26-3:         Summaryof field investigations quantity

Facility Quantity
Borehole Test Pit Geophysics lines Geomechanical Station
Tailings<br> Storage Facility (TSF) 9 15 5 2
Waste<br> Rock Storage Facility (WRSF) 5 10 5 2
Process<br> Plant 2 3 2 -
Water<br> Treatment Plant 2 2 2 -
Stockpile 3 6 6 -

Within the framework of the proposed geotechnical investigation program, the extracted core samples shall undergo both physical property testing and specialized tests, as determined by engineering judgment.

Samples must be transported using ASTM methods to ensure proper handling and consistency. Tailings samples should be obtained from a pilot plant, reflecting actual processing conditions. All samples must be representative of the deposit to provide reliable inputs for feasibility-level chemical studies. The required laboratory tests are illustrated in Table 26-4.

Table 26-4:         Summaryof Laboratory Tests

Test Material ASTM
Soil
Water<br> Content Foundation<br> and tailings ASTM<br> D2216
Atterberg<br> Limits Foundation<br> and tailings ASTM<br> D4319
Particle<br> Size Distribution Foundation<br> and tailings ASTM<br> C136
Specific<br> Gravity Foundation<br> and tailings ASTM<br> D854
One-dimensional<br> consolidation Foundation<br> and tailings ASTM<br> D2435
Consolidated<br> isotropic undrained (CIU)Triaxial Foundation<br> and tailings ASTM<br> D2850
Rheological<br> tests Tailings
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Test Material ASTM
--- --- ---
Rock
Unconfined<br> compressive strength of rock core specimens Foundation ASTM<br> D7012
Point<br> load strength test Foundation ASTM<br> D5731
Specific<br> gravity of rock Foundation ASTM<br> C97
Triaxial<br> compression test of rock specimens Foundation ASTM<br> D7012

Finally the development of both factual and Interpretative reports based on data gathered from the geotechnical field and laboratory programs.

The estimated cost for this work is $1.4 million.

26.7.2 TSF and WRSF Analyses and Studies

Due to the conceptual nature of this study and the limited information available at the time of writing, several assumptions have been made regarding the layout, MTOs, and construction of the proposed TSF and WRSF. Geotechnical properties of construction materials will be required to support slope stability analyses and other geotechnical assessments needed to confirm that the TSF and WRSF can be constructed as designed. Additionally, the future geochemical analyis of the tailings and waste rock to determine if considerations for acid generation and metal leaching need to be incorporated into the desgin of these facilities.

Additional studies and data collection will be required to advance project development beyond the conceptual stage. Some—though not necessarily all—of the current data gaps that should be addressed in future studies include the following:

● Perform<br> deterministic and probabilistic local seismic hazard study for the development of design<br> seismic for infrastructure.
● Collection<br> of hydrological information through site-specific climate and hydrological studies to support<br> the design of ponds, diversion channels, spillways, development of the TSF and WRSF water<br> balances and tying in with the site-wide water balance.
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● Collection<br> of hydrogeological information from desktop studies and site investigations to improve understanding<br> of subsurface flow regimes and their implications for facility design and performance.
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● Confirm<br> geochemical characterization of tailings and waste rock from additional waste characterization<br> studies.
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● Develop<br> seepage predictions and seepage control measures for the TSF and WRSF based on<br> geotechincal field and laboratory programs.
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● Update<br> the tailings and waste rock deposition strategy to optimize material handling.
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● The<br> stability model should be reviewed and updated, as required, with consideration of the final<br> deposition and stacking plan using updated data on the material properties of the waste rock,<br> tailings, and the foundations for the TSF and WRSF.
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● Perform<br> a liquefaction assessment of foundations, considering updated information on material properties<br> for TSF.
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● Perform<br> dam breach analysis.
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● Solicit<br> budgetary quotes for earthworks and geosynthetics (i.e., geomembrane, geotextile, and piping)<br> to get more accurate pricing for the next cost estimate.
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● Develop<br> material take-offs and cost estimates (i.e, capital, sustaining capital, and operating costs)<br> for TSF and WRSF.
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The additional information obtained and the assumptions made in this study can be verified or updated to advance the project to the next level of design. The cost of implementing the above recommendations for geotechnical engineering design is estimated at $0.6 million.

26.7.3 Hydrogeology

Hydrogeological drilling and monitoring are recommended to build a conceptual model that will later be refined into a numerical model. The recommended activities by stages are:

Stage1: Advanced Exploration/Pre-feasibility

● Preliminary<br> hydrogeological survey (mapping of springs, streams, existing wells).
● Installation<br> of exploratory piezometers in strategic zones.
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● Initial<br> measurement of groundwater levels and seasonal variations.
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● Preparation<br> of a conceptual hydrogeological model.
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● Preliminary<br> analytical estimate of pit groundwater inflow
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● Basic<br> hydrochemical sampling.
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Stage 2: Feasibility and Mine Design

● Dedicated<br> hydrogeological drillings (network of observation wells).
● Pumping<br> tests and slug tests to determine hydraulic parameters.
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● Recharge<br> studies and water balance.
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● Continuous<br> monitoring of water levels and quality.
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● Numerical<br> modeling of groundwater flow and prediction of open-pit impacts.
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● Definition<br> of drainage and pumping systems.
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● Evaluation<br> of interaction with surface water bodies.
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● Water<br> management plan (collection, treatment, reinjection, or discharge).
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The estimated cost forthis work is US$1.5 million for Stage 1 and US$2.0 million for Stage 2. Progressing to Stage 2 is not contingent on results from the Stage 1 work.

26.7.4 Hydrology

The hydrological work plan for engineering purposes is divided into four parts and must include the minimum described:

● Climatology<br> workplan:
o Statistical<br> analysis of climate data for the purposes of:
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§ Total<br> annual precipitation and evaporation for average, wet, and dry years
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§ PMP<br> analysis
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The estimated cost for this work is $1.0 million.

26.7.5 Geochemistry

The recommended geochemistry workplan comprises:

● Review<br> of latest available data and mine plan information, to include drill hole logs, resource<br> reporting, geological descriptions etc.
● Design<br> of sampling and analysis plan, to include:
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o Sample<br> selection in coordination with NPM geology team
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o Completion<br> of static and kinetic laboratory-based test programs
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o Estimated<br> 400-600 samples comprising overburden, drill core (waste rock/mill feed), metallurgical test<br> tailings and legacy mine wastes.
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● Reporting<br> – to include data analysis and interpretation and implications for mine waste management<br> during all phases of mine life.
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● Deliverables<br> will include:
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o Draft<br> static geochemistry report
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o Draft<br> interim memo on laboratory kinetic test results (to week 20)
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o Draft<br> and final reports on static and kinetic test programs (to week 40).
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The estimated cost for this work is US$2.5 million.

26.8 Environmental Studies

The following recommendations are made regarding future studies and activities related to areas of environment, permitting and community engagement. Subject matter experts should be retained to complete these studies and activities which will support the Project though the feasibility stages and provide a strong basis for feasibility design, future permitting, and community acceptance:

● Building<br> on the baseline studies completed to date, commence the design and implementation of a multi-year,<br> seasonal baseline environmental studies program to support the Environmental Impact Assessment<br> (EEIA) and environmental licensing process in accordance with Bolivia's Environmental Law<br> No. 1333, applicable MMAyA and the Ministry of Mining and Metallurgy (MMM) requirements,<br> and international best practice (i.e., IFC Performance Standards).
● To<br> facilitate the implementation of the above recommendation, complete a gap assessment of available<br> environmental and social baseline environmental data and reports relative to permitting,<br> feasibility design, and community requirements that can be used to design and prioritize<br> socio-environmental baseline updates and programs with the overall objective of reducing<br> permitting timelines. This assessment should focus on the gaps in studies identified to date<br> which include geochemistry and ARD/ML potential, noise, hydrogeology, faunal surveys, species<br> at risk identification for flora and fauna, aquatic and fish studies, archaeological and<br> cultural resource surveys, and focused studies on community land use within and adjacent<br> to the Project area based on primary data sources.
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● Based<br> on the assessment above, build upon and enhance the baseline studies and sampling/monitoring<br> network established for the 2024 baseline studies programs, adding new studies and additional<br> sampling/monitoring locations as required.
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● To<br> assist in the design and development of the Project, environmental constraint mapping should<br> be conducted and periodically updated, based on the results of historical and future baseline<br> environmental and land use/cultural studies. This mapping should be utilized to derisk the<br> Project at the early design stages of the Project.
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● Groundwater<br> and surface water studies should be implemented to support the development of an integrated<br> water balance model for the mine that would ensure sufficient process water availability,<br> model pit inflows and seepage, and predict effluent quality. It will be important to identify<br> and predict potential impacts, if any, to downstream surface and groundwater resources that<br> are important for the maintenance of nearby lotic and wetland ecosystems as well as agricultural/livestock<br> requirements for the local community. These costs are included under Section 26.9.
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● A<br> comprehensive geochemical testing program should be conducted which includes waste rock, tailings, ore stockpiles,<br> overburden and mill feed characterization, formal acid–base accounting, elemental analyses,<br> kinetic testwork, mineralogical studies, water-rock interaction modelling, geochemical source<br> term modelling, and associated data management, QA/QC, interpretations and reporting. The<br> exploration data base and model should be reviewed and used as a basis for developing a screening<br> level geochemical sampling program focused on available drill core and samples. Sample selection<br> should be based on an understanding of geology including lithotypes, mineralogy, alteration,<br> and structure for waste rock, and tailings. The screening level geochemistry assessment should<br> be based on mining industry standard guidance for geochemical characterization such as that<br> described by MEND (Price, 2009). The results of the screening level work should be used to<br> design a more comprehensive program involving a much large number of samples. These costs<br> are included under Section 26.9.
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● Continue<br> to engage with government, local groups, and communities to align the Project with external<br> expectations and requirements, and to develop open and transparent relationships to help<br> drive project success. To support ongoing engagement activities, consideration should be<br> given to designing a socio -economic and cultural environment management and monitoring program<br> that would help to track information on community conditions, local development, and the<br> interaction between project activities and social dynamics.
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● Commence<br> the application to Bolivia’s Ministry of Environment and Water to obtain its environmental<br> categorization as a proposed open pit operation and formally commence the EEIA process, required<br> technical work and EEIA document preparation.
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The total cost for this work is estimated at US$6.0 million.

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27 References
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ALS Kamloops. (2023a). Metallurgical Testwork on Composites from the Carangas Project. New Pacific Metals Corp, Project KM6848, May 31, 2023.

ALS Kamloops. (2023b). Comminution Testwork on Composites from the Carangas Project. New Pacific Metals Corp, Project KM7065, September 8, 2023.

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ALS Kamloops. (2025). Metallurgical Testwork on Composites from the Carangas Project to Support a Pre-Feasibility Study. New Pacific Metals Corp, Project KM7100, Final Report, August 14, 2025.

Arce Burgoa. (2009). Metalliferous ore deposits of Bolivia. Second edition. p.369.

Bedient, P. B., Huber, W. C., & Vieux, B. E. (2019). Hydrology and Floodplain Analysis (6th ed.). Pearson.

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CIM, (2014). CanadianInstitute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and Mineral Reserves. Adopted by the CIM Council on May 10, 2014.

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Instituto Nacional de Estadística (INE) (2012, 2014). – Census data National Population and Housing Census) and demographic/economic estimates.

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Price, W. A. (2009). Prediction manual for drainage chemistry from sulphidic geologic materials (MEND Report 1.20.1). MEND Program. https://mend-nedem.org/mend-report/prediction-manual-for-drainage-chemistry-from-sulphidic-geologic-materials/

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Sturm, T. W. (2001). Open channel hydraulics (International ed.). McGraw-Hill

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